INI S-mf—9098 ABSTRACTS of papers to be presented at

INTEK 50 ' fil f

International Conference on Recent Advances in Mineral Science and Technology

Sandton, 26th to 30th March, 1984 •J

j ABSTRACTS 1 1 —•• "1 of papers to be presented at

MINTEK 50

Published and printed by the Council for Mineral Technology (Mintek) 1984 CORRIGENDA ABSTRACTS OF PAPERS TO BE PRESENTED AT MINTEK 50

(1) Paper A 7/2 (page 53) Abstract not yet available.

(2) Paper C1/1 (page. 110) The title and abstract are as follows:

SESSION Cl Modern Steel Processes

Paper Cl/1 THK MRKCT USE OF COAL IN IRON- AND STEEL-MAKING OPERATIONS

by

L. Von Bogdandy, M. Chitil, J.A. Innes, and C.G. Jonker Kloeckner CRA, West Germany

Tlu* use of widely available low-cost bituminous coal in steelmaking to compete with coke-based or electricity-based processes was adopted by the Kloeckner-CRA joint venture as the basis of technology development in the steel industry. This led to the introduction of the Allothermic converter and steelmaking processes such as KMS and KS, which permit the partial and full replacement of hot metal by scrap in steelmaking. These processes feature the pneumatic injection of coal into an OBM converter, in which part of the coal energy is used for melting and part is converted into a gas of medium calorific value. This gas is used to replace natural gas or oil requirements in existing steel plants. Post-combustion of the gas in the converter itself is a major tool in increasing the proportion of available energy in coal for melting purposes. The control of post- rom')ustion is therefore an important area of technical innovation in steelmaking. The very flexible operating mode of the AHothermic converter permits the economically optimal combination of hot metal-scrap ratios, steel production, and gas production for each plant location. Industrial-scale pilot-plant work on the use of the AHothermic converter as a coal- gasification medium has been completed. Further development work coupling the AHothermic converter with a direct-reduction unit indicates great potential for the reduction of operating and investment costs in the steel industry. 7

CONTENTS

The letter and numbers preceding each title indicate the session in which that paper is to be presented, and the second number indicates the position of the paper in a parallel session, e.g. B2/6 means that the paper is the sixth presentation in Session B2. It should be noted that the titles and authors are not necessarily the same as they will appear in the final programme. They appear here as they were at the time of going | to press.

Plenary Sessions i Page PI Plasma Metallurgy in the 1980s, by K.J. Reid 1 ; P2 Recent Advances in the Leaching of Sulphides and in the Precipitation of Iron, byJ.E. Dutrizac 3 P3 Ore-dressing Developments in North America, by A.L. Mular 5 P4 Developments in Separation Science in Hydrometallurgy, by D.S. Flett 5 • P5 Recent Developments in the Modelling and Control of Mineral Processing | Plants, byAJ. Lynch and D.J. McKee 7 ' P6 Progress in Instrumental Analysis: Developments in Atomic-absorption, X-ray-fluorescence, and Plasma-emission Spectrometry for the Analysis : of Metals and Ores, by R.L. Watters, Jr 9

Session Al: Optimization of Ore-dressing Processes and Energy Saving Si A1/1 Improvement of Dynamic Dense-medium Processes by Using Two-stage f Separation Systems with Recirculation of Middlings, by G. Ferrara, H.J. ! Ruff, and G. Schena 11 A1/2 Fundamental Factors Affecting the Efficiency of Fine-coal Beneficiation: the Petrographk and Mineralogical Composition of Coal, by L.M. Falcon and R.M.S. Falcon 12 A1/3 Investigation of the Latest Crusher Technology at Rustenburg Platinum Mines Ltd—Union Section, by L.A. Cramer and R.E. Phillips 13 A1/4 An Applied Mineralogical Investigation into the Production of a Pyrite Concentrate from the Roan Antelope Ore Deposit, Zambia, by M. Pearl and N. Kostic 15

l Session A2: Optimization of Ore-dressing Processes and Energy Saving (continued) • A2/1 The Improvement of Hydrocycloning Efficiency by Modification of the i Vorticity Function of the Flow Field, by M.D. Brayshaw and E.T. j Woodburn 17 A2/2 Control and Optimization of the Operation of Milling Circuits, by D.G. Hulbert and I.J. Barker 19 A2/3 The Measurement of Parameters Describing the Dynamic Behaviour of the Load in a Grinding Mill, by M.H. Moys 20 A2/4 The Effects of Changes in Liner Design and Mill Speed on Rod-mill Parameters, by D.D. Howat and L.A. Vermeulen 21 Session A3: Optimization of Ore-dressing Processes and Energy Saving (continued) -•> A3/1 Flotation Machine Scale-up: the Path Towards Improved Process Economics, by V.R. Degner 23 A3/2 The Influence of Polymeric Depressants on the Adsorption of Thiol Collectors in Sulphide Flotation, by E. Sieenberg and P.J. Harris 25 A3/3 The Hydrophobic Character of Semi-soluble Salt Minerals with Oleate as Collector, byJ.D. Miller and M. Misra 26 A3/4 A Study of the Selective Flotation of Fluorite from Calcite by the Use of a Single Bubble-stream Microflotation Cell, by E.W. Giesekke and P.J. Harris 28 Session A4: Treatment of Fines, Tailings, and Low-grade Ores A4/1 Surface Interactions in Fine Particle Flotation, by P. Somasundaran. 29 A4/2 Pyrite Flotation with Diethyldithiophosphate and Mercaptobenzthiazole: an Electrochemical Study, by D.R. Groot 31 A4/3 Improvements in the Flotation of Low-grade Fluorspar Ores, by P.G. Kilian 32 A4/4 Gold Recovery by Flotation at Bougainville Copper Ltd, by N.C. Clarke and E.W. Beermann 33

Session A5: Treatment of Fines, Tailings, and Low-grade Ores (continued) A5/1 The Role of Waste Sorting in the South African Gold Mining Industry, by J.S. Freer and R.C. Bohme 36 A5/2 Selective Magnetic Flocculation of Fine, Weakly Magnetic Minerals, by J. Svoboda 37 A5/3 Single-wire Magnetic Separation: a Diagnostic Tool for Mineral Processing, byJ.H.P. Watson and D. Rassi 40 A5/4 Bacterial Heap Leaching of Low-grade Nickel Material, by P.C. Miller, I.J. Corrans, and A.J. Southwood 42

Session A6: Innovations A6/I The Design and Construction of a Superconducting Magnet for Production-scale Dry Separation of Minerals, by J.A. Good and K. White 44 A6/2 Industrial-scale Dry Beneflciation of Phosphate-bearing Pyroxenite Ore, by E.H. Roux, J.G. Goodey, E.F. Wepener, and K.R. Hodierne.... 45 A6/3 Fast Flotation with an Air-sparged Hydrocyclone, by J.D. Miller and D,J. Kinneberg 46 A6/4 Some Engineering Guidelines on Flotation Reagent Use, by R.R. Klimpel 49

Session A7: Modelling Design and Control A7/1 The Process Engineering of Size Reduction Using Ball Mills, by R.R. Klimpel and L.G. Austin 52 A7/2 Particle Trajectories and Charge Shapes in Centrifugal Milling, by D.I. Hoyer 53 J

A7/3 Multivariable Control at Frank Concentrator, by I.J. Barker, S.C. Axcell, and D.G. Hulbert 53 A 7/4 The Use of Simulation in the Operation and Design of Mineral Processing Plants, by D.J. McKee and A.J. Lynch 54

Session A8: Modelling Design and Control (continued) A8/1 Recent Advances in On-stream Analysis Equipment, by A. Toop, L.R. Wilkinson, and G.J. Wenk 57 A8/2 A Recent Advance in On-stream Analysis Equipment for Mineral Concentrators and Metal Refineries, by K. Saarhelo and D.R. Barker 59

Session Bl: Electrolytic Processes Bi/1 Recent Developments in the Design of Electrochemical Cells for the Recovery of Metals, by R. Kammel 61 Bl/2 The Electrodeposition of Manganese Dioxide: Theory and Practice, by R.L. Paul and A. Cartwright 62 Bl/3 The Electrodialytic Generation of Sulphuric Acid and Ammonia from Dissolved Aqueous Ammonium Sulphate, by P. Steel and B. Verbaan 63 Bl/4 The Electrowinning of Chromium from Trivalent Chloride Solutions, by M.D. Birkett 65

Session B2: Equipment and Modelling B2/1 Kinetic Models for the Adsorption of Gold onto Activated Carbon, by J.S.J. van Deventer 68 B2/2 The Measurement of Carbon Concentration in the Adsorption Vessels of a Carbon-in-pulp Plant, by N.D. Hulse, B. Fitzgerald, and M.J. Ohlson de Fine.. 69 B2/3 An Investigation into Some of the Factors Influencing the Filtration of Metallurgical Slurries, by G.F. Lahoud and W.A.M. te Riele 70 B2/4 NIMCIX Absorption Columns for the Recovery of Uranium at Vaal Reefs South, by M.A. Ford and R.J. Kelly 72

Session BS: Leaching Processes B3/1 Composition and Phase Changes during Oxidative Acid Leaching Reactions, by A.R. Burkin 73 B3/2 • Uranium-bearing Minerals in Witwatersrand Rocks, and Their Behaviour During Leaching, by G. Smits 74 B3/3 The Mode of Occurrence of Gold and Silver in Dominion Reef and Their Response to Cyanidation After Pressure Leaching, by G.W. Glatthaar and C.E. Feather 76 B3/4 Modelling of the Hartebeestfontein Gold Mine Uranium Leaching and Ion Exchange Processes and Its Role in Economic Plant Operation, by B.R. Broekman and B. Ward 77 Session B4: Leaching Processes (continued) B4/1 Laboratory Testing of UJudag Scheelite Concentrate for the Production of Ammonium Paratungstate, by Y.A. Topkaya and H. Eric 79 B4/2 The Direct Leaching of Oxidic Manganese Ores, by W.A. Nattrass and W.A.M. te Riele 80 B4/3 Enhanced Leaching Kinetics of Chalcopyrite from CuFeS2/C Paniculate Aggregates by Ferric Sulphate, by R.Y. Wan, J.D. Miller, and G. Simkovich 82 B4/4 The Leaching of Base Metals from the Calcines Produced by the Roasting of Pyrite Concentrates, by M.J. Nicol and A.O. Filmer 84

Session B5: Separation Processes B5/1 Precipitation of Metal Values from Cationic Extractants, by AJ. Monhemius 86 B5/2 The Separation of Chromium and Iron by the Solvent Extraction of Ferrochromium Leach Liquors, by J.S. Preston 38 B5/3 Zinc, Manganese, and Sulphuric Acid Recovery by Solvent Extraction from Spent Electrolyte, by D. Buttinelli, C. Giavarini, and A. Mercanti 89 B5/4 A Review of the Development of Resins for Use in Hydrometallurgy, by B.R. Green 91

Session B6: Separation Processes (continued) B6/1 A Resin-in-leach Process for the Extraction of Manganese from an Oxide, by M.W. Johns and A. Mehmet 93 B6/2 Studies on the Mechanism of Gold Adsorption on Carbon, by N. Tsuchida, M. Ruane, and D,M. Muir 94 B6/3 Practical Considerations of Gold Recovery by Carbon from Main and Effluent Streams at Vaal Reefs, by W.A. Kretschmer, R.F. Dewhirst, RJ. Human, R.J. Kelly, and B. Strong 95 B6/4 A Novel Application of Resin- in-pulp in the Metallurgical Industry, by C.A. Fleming and G. Cromberge 97

Session B7: Process Development

B7/1 Recent Advances in the Development of Hydrometallurgical Processes i / for the Treatment of Base Metal Sulphides, by E.D. Nogueira 99 B7/2 The Carbon-in-pulp Plant at Rand Mines Milling and Mining Company Limited: Problems Encountered and Developments Introduced, by P.A. Laxen and T.D. Brown 100 B7/3 The Lodeve Mill: a Complex Alkaline Process for a Complex Uranium Ore, by R. Bodu, J.P. Herbert, P. Lafforgue, G. Lyaudet, P. Michel, and R. Vanhelleputte 101 B7/4 The Production of Electrolytic Manganese Dioxide from Ferromanganese Furnace Sludge, by C.F.B. Coetzee and W.A.M. te Riele 103 J

• j Session B8: Process Development (continued) ! I B8/1 Recovery of Lead from Mixed Sulphide Concentrates, by A.O. Filmer \ $ and G.G. Briggs 106 j |; B8/2 An Alternative Route for the Production of Chromium Chemicals from I | Chromite, by M.F. Dawson and R.I. Edwards 108 : i Session Cl: Modern Steel Processes Cl/1 Economic and Technical Aspects of the Direct Injection of Coal in Iron- and Steel-making Operations, by L. Von Bogdandy, M. Chitil, J.A. Innes, and C.G. Jonker 110 Cl/2 The KR Process, a New Development in Steel Making, by E. Golde 110 Cl/3 The Behaviour of Sulphur in a Kiln for the Production of Sponge Iron, by J.A. Theron and R.J. Dippenaar Ill ; Cl/4 (Title and abstract not yet available) 112 Session C2: Furnace Development ! C2/1 The Operation, Control, and Design of Submerged-arc Ferro-alloy [ ^ Furnaces, by M.S. Rennie 113 C2/2 Operational Parameters for Stfderberg Electrodes from Calculations, Measurements and Plant Experience, by R. Innvaer, L. Olsen, and A. Vatland 114

..• Session C3: Plasma Technology I C3/1 Metallurgical Reaction Philosophies of Transferred-arc Plasma Furnaces, | by K.U. Maske and K.G. Reid 117 •• C3/2 The Dissipation of Energy in d.c. Transferred-arc Plasma Systems and the Consequences for the Production of Ferro-alloys, by A.B. Stewart 118

Session G4: Plasma Processes - C4/1 Non-ferrous Metals Extraction Using Tetronics Plasma System, with Particular Emphasis on Zinc and Lead, by D.G. Page and C.P. Heanley 120 C4/2 The Application of Transferred-arc Plasma to the Melting of Metal Fines, by L.B. McRae, N.A. Barcza, and T.R. Curr 121 C4/3 Industrial Processes Based Upon Plasma Technology, by S. Santen and J. Skogberg 123 • C4/4 Smelting Reduction of Composite CL»*omite Pellets by Plasma, by S. , Kouroki, K. Morita, and N. Sano 125 ! t i ' j Session C5: Plasma Facilities ; C5/1 The 3,2 MVA Plasma Facility at Mintek, by T.R. Curr, K.C. Nicol, J.F. Mooney, A.B. Stewart, and N.A. Barcza i28 C5/2 An Extended Arc Reactor, Using Graphite Electrodes, by C.B. Alcock, A. McLean, and I.D. Sommerville 129 C5/3 The Development of a High-power Arc Heater—A Progress Report, by A.L. Hare 130 I

! C5/4 Three-phase a.c. Plasma Melting: State of Development and Potential > Applications, by H.J. Bebber, J. Hartwig, D. Neuschutz, and H.-O. '• ~. Rossner 131

I Session C6: Mineral and Material Science in Pyrometallurgy C6/1 Slag-metal Equilibria Involving Chromium as a Component, by A. Muan 134 C6/2 The Mineralogy of the Reduction of Chromium Ore in a Plasma Furnace, by A. Wedepohl 135 C6/3 Chromite Mineralogy and Metallurgical Behaviour, by T.R.C. Fernandes 137 C6/4 From Metallurgical Grade Silicon to Solar Grade Silicon—the Chemistry of Different Approaches, by T.N. Lung 138

Session C7: Non-ferrous Pyrometallurgy i C7/1 Extraction of Cobalt from Nkana Smelter Converter Slags, by E. Moyo, ! J. Kalusa, and S.L. Kher 140 C7/2 Sprinkle Smelting Development, by T.D. Jackson and J.A. Schneider 142

Session C8: Applied Measurements C8/1 Factors Affecting the Corrosivity of Underground Mine Waters, by R.T. White and A. Higginson 143 C8/2 Continuous Monitoring of Information from the Inside of a Rotary Mill, by L.A. Vermeulen, M.J. Ohlson de Fine, and F. Schakowski 144

Session C9: Instrumental Analysis C9/1 Modern Trends in X-ray Powder Diffraction, by R.L. Snyder, H.E. Gobel, and G. Hubert 145 C9/2 Automatic On-line Analysis with Atomic-absorption Spectrophotometry, by R.V.D. Robert and G.T.W. Ormrod 147 C9/3 Image Analysis of Gold-bearing Ores and Products, by E.J. Oosthuyzen 148 C9/4 The Application of X-ray Fluorescence Spectrometry in Minerals Processing, byJ.P. Willis 149

Session CIO: Instrumental Analysis (continued) \ C10/1 Application of Ion Chromatography to Mineral Processing, by C. Pohlandt 151 1 C10/2 The Direct Analysis of Solid Materials in the Form of Slurries by the : Inductively Coupled Plasma Technique, by A.E. Watson and P. Humphries-Cuff 152 J

PLENARY SESSIONS Paper PI

PLASMA METALLURGY IN THE 1980s

by KJ. Reid University of Minnesota U.S.A. Despite a considerable history of research there are currently very few commercial operations using plasma in larger-scale process metallurgical applications. In recent years many workers have attempted to employ known plasma-generating devices in process metallurgy, and three major divisions can be observed: heat transfer to molten- bath systems, direct interaction of the plasma medium with solid feedstocks, and heat transfer via plasma-generated high-enthalpy gas. In the first category, much progress has been made in scrap melting and refining applications while relatively smaller advances have been made in metallurgical reduction and smelting applications. In the second category, research in the 1960s and 1970s sought to generate an expanded plasma in an effort to reduce the very high thermal, velocity, and viscosity gradients in a conventional plasma arc and thereby facilitate the entry of solids. To date this work has not led to any successful commercial applications. In the third category, a small-scale commercial plant to recover metallics from process dust was announced in 1983. Current projects in the major international centres of research in plasma metallurgy are reviewed in the paper. Details of activities in the U.S.S.R. are not readily available, but the Russians are well known for their excellent work in plasma-generator design and for applications in melting, refining, and speciality products. It is probable that the installed power for commercial plasma-metallurgy systems in the U.S.S.R. is larger than in the rest of the World. Japan has made major progress in plasma-melting systems for a variety of materials ranging from titanium alloys to radioactive wastes, and the GDR has a 30 MW scrap- melting plasma furnace in operation. In Sweden, the main thrust has been towards the reduction of metal oxide materials, and has led to the announcement of a commercial unit to apply the SKF Plasmadust process for metal recovery from steelplant wastes. In South Africa, a major effort has been mounted at the Council for Mineral Technology (Mintek) with the support of the South African ferro-alloy industry and the commercial application of an ASEA direct-current arc furnace at the Krugersdorp plant of MSA, begun in 1983. A plasma-melting application using the GDR technology is also in operation. In the U.K. a commercial plant is in operation at Tioxide Ltd, and considerable research has been carried out by Tetronics on applications of the expanded precessive plasma (EPP) system to a range of metallurgical processes. Canada has shown considerable growth in plasma-metallurgy research in recent years, with fundamental work at Sherbrook and McGill, and applied research at Toronto, Noranda, and the Quebec Hydro Authority. In the U.S.A., small commercial applications are in operation for the treatment of zirconia, spheroidization of magnetite, partial remelt of titanium scrap, scrap melting, and electronic scrap recovery. There is little plasma research in process metallurgy at the universities, with the main activities confined to Columbia and Minnesota. The University of Minnesota has two plasma programmes: one in the Center for Plasma Chemistry (CPC) with people from several departments working in a range of plasma fields, and another, focused on process-metallurgical applications, being carried out in the Mineral Resources Research Center (MRRC). The CPC work covers low-pressure discharges, arc-attachment phenomena, heat transfer, and organic and inorganic chemistry in low- and medium-pressure plasmas. The work at MRRC is based on an expanded plasma system, termed a sustained- shockwave plasma (SSP), with investigations covering the reduction of taconite concentrate, treatment of copper-nickel concentrates, chromite reduction, manufacture of novel cements, and processing of coal wastes. Several reactors are available for process- metallurgical research with power sources available in the operating ranges 1 to 15 kW, 10 to 100 kW, 50 to 250 kW, and 200 to HOOkW. Major areas of interest are the non-equilibrium characteristics of the SSP medium, reactions far from equilibrium, and the kinetics of metallurgical reactions under SSP conditions. Important fundamental areas relating to plasma reactors include the generation of the plasma medium, torch life, heat and mass transfer, control of the chemical potential, and solid-plasma interactions, including the mechanics of feed systems and reactor geometry. In process metallurgy there is a continuing drive to seek improved efficiency, superior products, and reduced costs. The high cost of electrical power derived from fossil fuel favours plasma applications for high-value products and for materials that currently use electric arc-furnace technology. The main technical problems to be solved lie in the areas of reliability, life, and scale-up of plasma generators, and in optimizing plasma-reactor geometry for specific metallurgical applications. Government and private funding for research in plasma metallurgy has not been very generous in the past decade, and consequently only modest progress has been made. However, recent economic problems have served to focus attention on the need for increased research activities in the basic industries, and major advances in research and in the commercial demonstration of plasma systems for process metallurgy are expected in the remainder of the 1980s. 7

Paper P2

RECENT ADVANCES IN THE LEACHING OF SULPHIDES AND IN THE PRECIPITATION OF IRON

by

J.E. Dutrizac CANMET Canada Significant advances have been made in many areas of sulphide hydrometallurgy; progress has tended to be more evolutionary than revolutionary, despite early claims to the contrary. Although hydrometallurgy avoids sulphur dioxide emissions, it has its own environmental problems and is often energy-intensive. Within such a general framework, this paper examines some of the recent advances, both in plant practice and in process fundamentals, which have occurred in the leaching of sulphides and in the closely associated area of iron precipitation. Oxygen-Sulphuric acid pressure-leaching techniques have been successfully extended to zinc concentrates. This process dissolves zinc concentrates directly at 150 °C and 620 kPa of oxygen pressure to form soluble zinc sulphate and elemental sulphur. Over 98 per cent zinc is solubilized from conventional concentrates; pyrite is only superficially leached. Advantages are the elimination of the roaster circuit, gas-handling plant, and ferrite-treatment operations, as well as the inherent decoupling of zinc production from sulphuric acid manufacture. The key role of surfactants in achieving high recoveries is discussed. Problems still remain with chloride-fluoride control and with lead-silver losses as plumbojarosite. Pressure-leaching techniques have been extended to copper concentrates, to refractory gold ores, and to uranium ores with major arsenide mineralization. Pressure-leaching processes for cobalt sulphide feeds have been developed but not yet implemented. Recent work on oxygen-sulphuric acid pressure leaching in the presence of chloride could extend the range of applications of this technology. Progress continues to be made in both the practice and the theory of copper- concentrate leaching in ferric chloride and/or cupric chloride media. The Duval process has advanced to the demonstration-plant scale; high copper recoveries with elemental- sulphur formation have been achieved. Several other chloride-leaching processes (e.g. Cymet, Elkem, etc.) have been developed and tested at either the mini- or pilot-plant level. Most attempt to utilize the energy advantages of copper electrowinning from the Cu+' state to yield either easily detached crystals or compact plates. Work on silver recovery from FeCl2-CuCl media continues, but this has proved to be a difficult problem. Significant improvements have been made in the understanding of the factors affecting chalcopyrite leaching in chloride media, but many areas of uncertainty still remain. There is agreement that the initial stage of leaching involves the formation of an extremely thin protective layer on the chalcopyrite that is then covered with other layers, which may or may not be protective. The numerous studies related to both the chemical and electrochemical leaching of chalcopyrite are reviewed, and areas of consensus or disagreement are outlined. Recent developments concerning the role of silver catalysis in the leaching of chalcopyrite in both conventional and bacteriological systems are discussed. The hydrometallurgical processing of lead concentrates offers advantages from both the technological and the environmental points of view. The numerous (U.S. Bureau of Mines, Minemet, Reynolds, etc.) mini- or pilot-scale studk son lead hydrometallurgy are described, and the various metal-recovery options arc compared. Although significant progress has been made, problems remain with impurity control, silver recovery, and electrolytic-cell design for either aqueous or fused-salt systems. Comparatively little is known about the fundamentals of galena leaching, and the behaviour of the common silver minerals is largely an unknown field. Chloride-leaching processes have been extended to zinc concentrates and to pyritic zinc-lead-copper-silver concentrates. The Zincex process is a notable development, but many other pilot-scale investigations have taken place, especially for pyritic bulk concentrates. A major advantage for the chloride processing of such feeds is that pyrite is not attacked and most of the sulphur reports in the elemental form. In addition, several fundamental studies on the behaviour of zinc sulphides have been done, and these are assessed. Iron-precipitation processes, especially the jarosite process, have become sophisticated unit operations capable of being integrated into a variety of hydrometallurgical flowsheets. From a base in zinc processing, jarosite precipitation is being developed for copper and cobalt circuits. Various modifications have been made either to improve the jarosite-precipitation process or to yield a cleaner jarosite residue low both in valuable metals and in environmentally undesirable elements. The use of limestone, either directly or indirectly, has been developed for neutralization of the hydrolysis acid, and the low- contaminant jarosite process offers the potential of avoiding entirely the loss of neutralizing agent. Advances have been made in the understanding of jarosite precipitation, and of the nature and extent of impurity element losses into such compounds. Both the hematite and goethite processes are used in industry, but fewer developments have occurred for these technologies. Use of the hematite process is expanding, however, and methods for recovering lead and silver from such circuits are being developed. Major research is still needed to find uses and/or better impoundment methods for iron residues.

4 ,;<•'

Paper P3

ORE-DRESSING DEVELOPMENTS IN NORTH AMERICA

by

A.L. Mular University of British Columbia Canada Canadian and American ore-dressing advances and innovations since 1976 in research and development, plant design, plant operation, and plant automation are reviewed. The major goals have been to increase efficiencies and reduce costs.

Paper P4

DEVELOPMENTS IN SEPARATION SCIENCE IN HYDROMETALLURGY

by

D.S. Flett Warren Spring Laboratory England

Hydrometallurgy in extractive metallurgy dates back several hundred years. For a long time, however, applications were limited, and only impure materials were recovered by precipitation processes such as cementation for copper. To quote Professor Wadsworth, 'modern large scale hydrometallurgical processing had to wait for the introduction of electrical power needed for electrolytic reduction and power for transport and agitation of large tonnages of ore slurries and solutions'. With the development of sophisticated separation processes for the concentration and purification of metal ions from dilute and/or multimetal leach liquors, and the ability to interface these with the final metal-winning step, it became possible to produce pure metals by hydrometallurgical means from low-grade and complex materials. In this context, it is the major unit processes of ion exchange, solvent extraction, activated-carbon adsorption and, recently, membrane processes that have had the greater impact on hydrometallurgy in recent times. Ion exchange played a major part in the development of the uranium industry only to see fixed-bed processes replaced by direct solvent extraction, which in turn gave way, at least in part, to continuous ion exchange. Ion exchange has never really made a large impact throughout hydrometallurgy; unlike solvent extraction, which has turned out to be as important a development in hydrometallurgy as froth flotation was for mineral processing. Nowadays, however, new chelating ion exchangers and resin impregnates are available which are showing signs of adoption in commercial plants. Carbon adsorption for gold, in particular the recently developed carbon-in-pulp processes, has given considerable flexibility to the type and grade of gold ore treatable by cyanidation and more efficient gold recovery. Carbon adsorption has permitted the development of heap leaching of low-grade gold ores, while carbon-in-pulp has materially reduced solid-liquid separation costs and has also permitted considerable reworking of old mine dumps in South Africa. The ideal technique for these applications would be ion exchange, and it may not be long before resins and processes are available for such duties. With solvent extraction now a mature process in hydrometallurgy, attention is turning to liquid membranes, which are now being heavily researched. Currently, however, there have not yet been any commercial applications. There are two types of liquid membranes, i.e. supported and emulsion membranes, which offer two different sets of advantages. Supported membranes provide a low-energy, low-cost continuous ion-exchange process. Their main drawback is the slow reaction rate, making applications difficult for depletion of metals down to very low concentration in, for example, effluents. Because the flux is concentration-dependent at low concentrations of metals, feed and bleed and similar back-mix reactors are not appropriate and plug-flow reactors are indicated. Emulsion membranes offer the advantages of multi-stage solvent extraction without the need for a multiplicity of stages. On the face of it, they should be ideal for effluent treatment, but unfortunately display the inherent disadvantages of solvent extraction in this application, i.e. soluble losses. Thus, the main application for supported liquid membranes is likely to be in main- stream hydrometallurgy, where liquors are returned to leaching although their use for metal extraction from unclarified liquors or pulps has considerable attraction. Exciting developments in the use of both liquid and ion-exchange membranes can be envisaged in hydrometallurgy. In the future, the use of volatile reagents is an intriguing possibility. Such an approach has already been exploited in analytical chemistry, using volatile metal diketonates in chromatography. Application to sulphide ores would produce hydrogen sulphide and the volatile diketonate of the metal of interest. Fairly high selectivities could result from a combination of complex stability constant and degree of volatility. In many ways, such a process would be no different from chloride volatilization, but much less corrosive! Finally, there is clearly a wider interest developing in non-aqueous chemistry in extractive metallurgy in general. The possibilities for the variation of ion activities in mixed aqueous-organic systems and in polar organic solvents are very wide indeed, leading to marked changes in ion-exchange behaviour and novel separation possibilities. The development of solvent-extraction systems with polar-non-polar immiscible organic solvents is entirely possible and could be of considerable interest in the longer term. This paper reviews the state-of-the-art of the unit processes outlined above. Paper P5

RECENT DEVELOPMENTS IN THE MODELLING AND CONTROL OF MINERAL PROCESSING PLANTS

by

A.J. Lynch and D.J. McKee Julius Kruttschnitt Mineral Research Centre Australia Over the past twenty years, a major world-wide research effort has been devoted to the mathematical modelling of many of the processes used in mineral dressing plants. A related development has occurred with the introduction of automatic control systems for many mineral processing operations. The modelling and control work have been strongly influenced by the increasing availability of powerful, low-cost digital computers. The other critical hardware item has been the on-stream analyser unit for on-line measurement of metal concentrations in flotation pulps. A third major factor in the development of modelling and control technology has been the realization by both research and operating personnel that, by the intelligent use of modelling and control techniques, very substantial gains can be achieved in the metallurgical and economic performance of production plants. It is not the intention in this paper to review the historical development of mathematical models and control systems used in mineral processing operations. The pioneering work in ball mill, hydrocyclone, and flotation modelling has been described many times. The early control systems for grinding circuits are well known, as are most of the first attempts at flotation control. After twenty years of effort, it is ciear that major progress has been achieved towards the goal of developing comprehensive models of a range of mineral dressing unit processes. However, it is also apparent that there are gaps in the existing knowledge. It is also asserted that the situation with regard to control system application for these same processes is not as advanced as is the case with mathematical models. The current status of models for comminution processes is briefly reviewed. While mathematical descriptions of crushers, and rod and ball mills are quite adequate, the same is not the case for autogenous and semi-autogenous grinding. One of the major problems in autogenous mill modelling is a lack of knowledge of the mill lead. This is a critical shortcoming, as the mill load has such a major influence on mill performance. The widely used empirical models of screens and hydrocyclones are adequate for simulation purposes provided their use is restricted to the data range from which they were developed. Work on developing more mechanistic classification models is continuing and this will ultimately result in more flexible classification models. Over the past few years, there has been an increasing application of comminution and classification models in the analysis of circuit design problems. The mineral processing industry is aware that computer simulation with adequate process models is a very powerful technique and can greatly reduce the effort involved in producing designs, as well as increasing the accuracy of the designs. The same techniques of modelling and simulation are being applied to an analysis of scale-up problems in ball mill design and to the ultimate application where a grinding circuit design is based on simple laboratory breakage tests on the ore and followed by simulation predictions. Models of the flotation process have been under development for twenty years also. These models are capable of predicting flotation circuit performance under restricted conditions. However, the very large number of variables which influence the flotation process greatly complicate the development of a comprehensive model. One of the major difficulties in flotation modelling has been an inability to adequately quantify the behaviour of composite particles. This difficulty has in turn been directly related to the inability to measure in an accurate way the degree of compositeness of particles. The problem has largly been solved with the introduction of automatic image analysers capable of providing quantitative information on the mineral composition of particles. It now seems that flotation modelling, which has not progressed greatly over the past five years, is in a position to again move forward. Automatic control system development for ball mill-classifier circuits has followed a steady and successful path. The new developments in grinding circuit control are associated with the introduction of modern control techniques which show promise in being able to refine the quality of control system performance. Control of autogenous milling circuits has proved difficult, and work is continuing. This is undoubtedly a critical area, in view of the increasing numbers of new autogenous grinding circuits. Automatic control of flotation circuits has proved to be a very difficult and elusive goal. When on-stream analysis became a reality in the early 1970s, there was an initial rapid attempt to install flotation control systems. However, ore type variability, which is present to some degree in all plants, has caused flotation control systems many problems. The problem of identifying the ore type before flotation commences has not yet been solved, and in recent years the trend in flotation control has been to simplify the strategies and the objectives and gradually build experience with control on the particular circuit. The pattern for the application of modelling and control in mineral processing is clear. Simulation techniques will become increasingly widespread and will be commonplace for circuit design problems. Control system development will be applied on a circuit- by-circuit basis using general concepts, but always recognizing the particular characteristics which make each control system an individual application.

8 Paper P6

PROGRESS IN INSTRUMENTAL ANALYSIS: DEVELOPMENTS IN ATOMIC-ABSORPTION, X-RAY-FLUORESCENCE, AND PLASMA-EMISSION SPECTROMETRY FOR THE ANALYSIS OF METALS AND ORES

by

R.L. Watters, Jr Center for Analytical Chemistry, National Bureau of Standards U.S.A. The routine analysis of metals and ores has traditionally involved either wet-chemical or instrumental spectrometric techniques. Instrumental analysis using atomic-absorption (AA), X-ray-fluorescence (XRF), and inductively coupled plasma-emission (ICP) techniques has gradually replaced many of the classical techniques over the past decade. This trend is a result of two developments in inorganic analysis. On the one hand, AA, XRF, and ICP techniques have improved to the point where they can achieve the precision and accuracy required for many analytical applications. On the other hand, technical personnel with expertise in wet-chemical separations and gravimetry are retiring from the analytical community without replacement, and the cost of non-instrumental analysis, which is often time-consuming, has become prohibitive. Some of the more recent developments in the techniques mentioned are reviewed, and ways in which the combination of classical and instrumental approaches can be used to advantage are discussed. When considering the history of instrumental analysis, one might assume that progress in well-established techniques like AA and XRF would be confined to application refinements rather than fundamental advances. Since the ICP is a relatively new technique, one would expect the more dramatic and fundamental developments to occur in this field. This is, however, not the case. Each of the techniques mentioned has benefited from both advancing technology in electronics and the vast experience of analysts in applications research. The availability of dedicated microprocessors in analytical instrumentation, coupled with this experience, has allowed the imagination of scientists and engineers to range over wide areas of instrumental improvements. As a result, physical, chemical, and spectroscopic phenomena can be precisely controlled to increase the accuracy and precision of routine analyses. Atomic-absorption spectrometry with electrothermal atomization (ETAAS) has been in use for over twenty years and has established itself as one of the most sensitive methods for determining trace metals. From its inception, the successful ETAAS instrument has had to deal with the inherently transient nature of the signal and the need for highly reproducible sample injection. Of interest here are the later developments which improved the accuracy of the net absorption measurement and reduced matrix effects. In complex matrices, such as those presented by metals and ores, the atomization of the analyte after the temperature of the furnace has stabilized at a predetermined level is an important key to accurate analysis. The proper use of matrix-modification techniques and the L'vov furnace with the stabilized temperature platform has achieved this condition. Another consequence of analysing complex matrices is the occurrence of structured and shifted spectral background in the region of the absorption line of interest. Continuum-source background correction has been only partially successful at ensuring the accurate measurement of net absorption in ETA AS. Recent advances in background-correction techniques using the Zeeman effect and a pulsed atomic-line source are promising approaches for accurate ETAAS analysis. While XRF is one of the most useful techniques for the precise analysis of solid materials, problems related to matrix effects can still impede the attainment of high accuracy. Progress in quantitative XRF at the National Bureau of Standards (NBS) has been achieved in the calibration of X-ray tube spectral output and the production of thin-glass films as standard reference materials (SRMs). The calibration efibrt involves the development of an algorithm to calculate the spectral distribution of X-ray tubes operating from 10 to 40 kV, with targets over a wide range of atomic numbers. These data are needed for fundamental parameter models which correct for absorption and enhancement effects in thick samples. Results comparing calculated line to continuum ratios with experimental values indicate that the NBS algorithm will provide more accurate data for such computer programmes. The process of certifying thin-glass films as SRMs is in progress. Twelve elements have been ion-beam sputtered onto two sets of films. The homogeneity, stability, and composition of the glass films are being measured using XRF, ICP, and neutron activation. Plasma-emission techniques like the ICP and the direct-current plasma (DCP) are the most recently developed for inorganic analysis. Both are capable of simultaneous multi-element analysis over a wide concentration range. The ICP is relatively free from matrix effects and, therefore, is a particularly attractive tool for the analysis of metals and ores. Commercial instrumentation has taken advantage of microprocessor control of data acquisition and handling. In particular, computer control of sequential wavelength spectrometers offers a degree of flexibility in spectral-line selection which is not available with polychromators. Some of the advantages, as well as possible compromises, in this approach are discussed. One of the problems in analysing metals and ores using the ICP is the conversion of the bulk sample to a form amenable to interference-free analysis. This has proved to be a difficult problem fur classical techniques, as well as instrumental spectrometry. The work at NBS has involved the examination of the spark discharge as a direct sampling device for the ICP. The preliminary results encouraged the commercialization of the spark-ICP combination, and studies of the mechanisms of solid sampling for accurate spectrometry continue. Finally, the use of an ICP as an ion source for mass spectrometry is described. Each unit provides the other with a significant advantage. Mass spectra are inherently less complex than optical-emission spectra, and the ICP can afford a convenient and efficient source of ions. Published data and applications are as yet scarce, but indications of analytical promise are discussed.

10 SESSION Al Optimization of Ore-dressing Processes and Energy Saving 1 Paper Al/1 IMPROVEMENT OF DYNAMIC DENSE-MEDIUM PROCESSES BY USING TWO-STAGE SEPARATION SYSTEMS WITH A RECIRCULATION OF MIDDLINGS h by G. Ferrara*, H.J. Rufff, and G. Schena* * University of Trieste, Italy TInpromin Ltd, England i ••"•• The improvements which can be achieved by using two-stage dense-medium separation systems with or without separate treatment or recirculation of middlings are reviewed, and both the technical and the economic aspects are discussed. In particular it is shown how the amount of coarse material which can be rejected, prior to more expensive processes, can be increased by feeding a two-stage dense-medium system with a relatively coarse top size and subsequently crushing only the small amount of middlings for proper liberation. It is then recalled that up to now only static three-product separators have been available to apply this concept in a single unit. However, as static separators cannot efficiently treat material much below 5 and 8 mm, this restricted the proportion of the run-of-mine ore which could be treated, and even more so the amount of middlings remaining above that size range after crushing for liberation. In contrast, dynamic systems can treat material efficiently down to between 0,5 and 1 mm, and with modern medium circuits even down to about 0,15 mm. Thus, they can treat a greaiter proportion of the raw ore and, by operating in a size range with inherently better liberation, can reject more coarse material, producing a higher feed grade for subsequent treatment and reducing the load on the water-reclamation system. So far the application of the two-stage approach has been discouraged because no dynamic three-product vessel has been available, and the extra cost and space required for two two-product plants in series has therefore been considerable. The paper describes the new three-product Tri-Flo dynamic dense-medium separator, which incorporates two stages in a single vessel, and which has proved capable of solving a number of difficult problems with good results and with insignificant extra investment or operation costs. To date the Tri-Flo Separator is operating in nine plants in Germany, Austria, Yugoslavia, Italy, Greece, and the U.K. The paper gives a short description of some of these plants, and reports on various situations where the recirculation or retreatment of middlings can be shown to be economically attractive for both coal and minerals. A method of computer simulation and prediction of industrial results that takes account of the possible improvement of the washability curve of the middlings by size reduction is described. This method allows the assessment and evaluation of both the technical and the economic factors relevant to the two-stage DMS process using different

11 separation densities and probable errors (2?p) for a variety of flowsheet configurations. (I Various flowsheet designs based on the retreatment of middlings, either by recirculation in the same separator or in other units, are compared and discussed. An example based on the Sardinian-Sulcis-Coal is presented, and shows that this technique can be particularly effective when used on difficult coals and ores.

Paper Al/2

FUNDAMENTAL FACTORS AFFECTING THE EFFICIENCY OF FINE-COAL BENEFICIATION: THE PETROGRAPHIC AND MINERALOGICAL COMPOSITION OF COAL

by

L.M. Falcon* and R.M.S. Falconf • Gold Fields of South Africa t Falcon Research Laboratories South Africa

With the production of higher proportions of fine coal (smaller than 0,5 mm) in modern mining methods, it has become necessary to investigate the various methods of preparation in order (1) to optimize the utilization of this size fraction and (2) to upgrade the products for the benefit of a number of selective markets. For this purpose, organic (maceral) and inorganic (mineral) composition is likely to become as important as calorific value, ash, volatile matter, and carbon contents. Previous work has shown that the beneficiation behaviour of fine coals may differ considerably in the various methods of preparation. Medium-rank, reactive-rich coals are known to float in froth-flotation processes more easily than low-rank (bituminous), inert-rich coals. Also, low-rank coals have been found to respond better to beneficiation by means of heavy-medium separation than by tabling or froth flotation. The purposes of the current paper therefore are (1) to choose two specific coal types typical of the low-rank Witbank-Highveld coalfield, (2) to investigate the organic and inorganic composition of the coals, and (3) to compare the behaviour of these coals and the distribution patterns of the components in heavy-medium separation and froth flotation. The two coal types selected for investigation were a bright, vitrinite-rich, high-volatile, low-ash coal (B) and a dull, inert-rich, low-volatile, high-ash coal (D). Both were subjected to heavy-medium separation (six floats and one sink each) and froth flotation (all the operational parameters being similar, followed by variations in the method of addition of the froth-flotation reagent). Full analyses of the original raw coals and the washed fractions were undertaken. These included proximate analyses (for ash, volatile matter, fixed-carbon, and moisture content) and petrographic analyses (maceral, mineral, group, microlithotype, and rank analyses). Additional analyses concerning mineral type, relative size, and form and degree of intergrowth were undertaken.

12 The results indicate that the coals are identical in rank, but vary very significantly in maceral, mineral, and microlithotype proportions. During the course of heavy- medium separation, both coals were found to possess very similar patterns in their distribution of organic and inorganic matter (i.e. quality), although the yields at different relative densities varied significantly. This implies good liberation based upon gravity, despite original composition. In froth flotation, however, the distribution patterns of the organic macerate and minerals varied significantly, both between the dull and the bright coals in identical operational procedures and between the same coals with varying patterns of flotation addition. These results indicate a complex situation relative to the surface chemistry of the organic and inorganic components of the coals, the varying proportions of the reactive to inert macerals, the varying types and relative size distributions of the minerals and of the degree of intergrowth of the mineral- maceral associations (tnicrolithotypes) typical of each coal, and the mechanical effect of the stepped addition of the froth-flotation reagent. Improvements in yield and distribution patterns were apparent under certain conditions. These results apply only to coals of low rank and typical Witbank composition. Once the fundamental compositional aspects of a coal are understood in relation to liberation behaviour, it will become possible to select and predict the optimum method of beneficiation for a variety of coals in a number of different geographical regions. Further problems relative to types and degrees of oxidation in coal and their effects on beneficiation may also be better understood.

Paper Al/3

INVESTIGATION OF THE LATEST CRUSHER TECHNOLOGY AT RUSTENBURG PLATINUM MINES LTD—UNION SECTION

by

L.A. Cramer and R.E. Phillips Rustenburg Platinum Mines Ltd South Africa

Union Section of the Rustenburg Platinum Mines is located in the northern Transvaal Bushveld, and has been in operation since the 1930s. Mortimer Concentrator, commissioned in 1973, was the focus of the investigation. The Mortimer crushing circuit consists of three stages; the second stage is preceded by wet screening, and the third stage is operated in closed circuit. A mass balance carried out on the crusher plant in July 1982 indicated that the plant was not being run at maximum power inputs. Circulating loads were high at approximately 200 per cent, crusher power draws were low, and the final product was coarse with the generation of only 9 to 13 per cent material smaller than 2 mm. Recent studies on crushing circuits around the world implied that significant improvements can be made in the plant's production efficiency. One of the new crushing technologies available in today's market is embodied in

13 the high-power-input, automatically controlled Allis-Chalmers hydrocone. Personnel of Johannesburg Consolidated Investments were at that time examining various alternatives for improving the Amandelbult crushing operation; for ease of comparison, Bateman's Engineering supplied a 3 ft hydrocone for testing as a tertiary crusher at Mortimer concentrator, where Nordberg cone crushers were in operation. The test programme was designed to accomplish the six tasks listed below: (1) establish the maximum power rate which can be applied to the Union Section ore and the optimum power rate within the machine's capabilities; (2) collect sufficient data at the optimum power rate to have statistical confidence in the operating results; (3) test the hypothesis that the Allis-Chalmers' Automatic Setting Regulation equipment and the hydrocone closed-side setting mechanism enable the hydrocone to operate continuously at very close to the maximum rated power draw; (4) test the hypothesis that higher power rates in crushing give rise to a finer crushed product, thus resulting in reduced circulating loads and a finer mill feed for the same closed-circuit screen size; (5) test the hypothesis that this greater amount of work in the crushing section will reduce the overall work required in the milling circuit; and (6) access the correlation between actual crusher performance and the prediction from twin-pendulum tests of the Allis-Chalmers Advance Technology Centre. All the above tasks were completed, but the effect of finer mill feed on the final mill product was not as marked as that predicted by the theory of the Bond Work Index. Finer crushing gave rise to a measurable amount of oversize in the flotation section, but the reduction in size of mil! product was considerably less than that predicted by calculations of the power input. The second most-recent crusher technology available on the market is reflected in the Omnicone 1560 crusher, manufactured by Rexnord. A second test programme involved a four-month trial of the Omnicone as a replacement for one of the existing 4>/4ft Nordberg short-head crushers. Thorough testing demonstrated that the Cnnicone can achieve the high power rates and fine crusher product produced by the Allis- Chalmers machine. The final phase of these investigations centred on the performance of the existing Nordberg tertiary circuit. The use of different liners was investigated, as were higher power rates than the traditional operation. Power draws can be increased to 70 per cent of the installed horse power and circulating loads reduced to 50 per cent, and similar amounts of fine material are present in the crusher product. Higher power inputs to the Omnicone and Nordberg short-head crushers did not, in the short testing period, give rise to any noticeable mechanical damage to the crushers. A much larger phase of testwork will be required to evaluate this area of concern. Plant practice at Bougainville and Rdssing indicates that the traditional Nordberg machines are not designed to operate at the consistently higher power rates. In conclusion, all the tested crushers can be made to produce equally fine products at equally high power rates. The mechanical reliability of various machines at the higher power rates has not been fully investigated at this time.

14 Paper A1/4

AN APPLIED MINERALOGICAL INVESTIGATION INTO THE PRODUCTION OF A PYRITE CONCENTRATE FROM THE ROAN ANTELOPE ORE DEPOSIT, ZAMBIA

by

M. Pearl and N. Kostic Zambia Consolidated Copper Mine Zambia

Pyrite is currently used on the Copperbelt of Zambia for the production of sulphuric acid. An increasing demand for acid has promoted an investigation into the production of a separate pyrite concentrate as a byproduct of the existing copper-flotation circuit at Luanshya. The ore from the Roan Antelope deposit, which provides the feed to the concentrator, consists predominantly of chalcopyrite and pyrite, with subordinate bornite and chalcocite associated with a dolomitic argillite. At present, the copper-bearing minerals are floated to produce a copper concentrate. Pyrite is also allowed to float to this concentrate, despite its non-economic value, as it aids matte formation during later smelting. In-plant testwork is being carried out to produce separate copper and pyrite concentrates by differential flotation. This is achieved by altering the collector and pH conditions and by using a pyrite depressant during the flotation of the copper sulphides. The first stage of this project involved testing the feasibility of obtaining separate concentrates by means of differential flotation. The next stage involved the production of rougher concentrates under more controlled conditions. A further stage is planned in the near future on the cleaning of the rougher concentrates. As an aid to the optimization of the process, and as a reference for the future cleaning testwork, various products were mineralogically assessed. This mineralogical study took the following form: (1) microscopic examination of the products from the preliminary tests in order to quantitatively assess the mineral distribution so that the validity of the project could be confirmed; (2) more detailed microscopic examination and image analysis of the products from the controlled runs, which included a mass-percentage distribution of the sulphide minerals, their liberation and degree of interlocking with one another, and the grain sizes of the sulphide minerals. Under controlled conditions and with a material balance for the feed, coppsr rougher concentrate, pyrite rougher concentrate, and final tailings, it was found that the recovery of copper sulphides to the copper rougher concentrate varies between 83 and 92 per cent, and to the pyrite rougher concentrate beween 2 and 8 per cent. Of the free grains, the optimum recovery to the copper rougher concentrate occurred in the 14 to 100/im size range. The larger free grains floated in the .pyrite rougher concentrate.

15 /•'

The majority of the locked copper sulphides were floating either to the copper rougher concentrate, or were lost to the final tailings, with very few being recovered in the pyrite rougher concentrate. The locked copper sulphides in the copper rougher concentrate were mostly larger than 50/im and, conversely, the majority lost to the final tailings were smaller than 50 pm. In all cases, most of the locked particles occurred in low- grade middlings. The results of the tests for pyrite were more variable than for the copper sulphides. This is probably caused by a large variance in the pyrite content of the feeds. Recoveries of pyrite to the pyrite rougher concentrate varied between 67 and 86 per cent, and to the copper rougher concentrate between 2 and 4 per cent. Very little of the locked pyrite was recovered by either rougher concentrate. These locked particles occurred mainly in low-grade middlings. The mineralogicai conclusions from this study should influence the progression of the testwork, particularly in the cleaning of the two rougher concentrates. As the liberation of the minerals is very high for both copper sulphides and pyrite in the concentrates, and the degree of association between the two mineral types is very low, there should be no reason why successful cleaning of the two rougher concentrates cannot be achieved. The coarse copper sulphides and coarse middlings reporting to the copper scavenger concentrate and the pyrite rougher concentrate indicate that these grains will not readily float in the copper rougher concentrate. However, regrinding the scavenger concentrate before it is returned to the head of the circuit should increase the recovery of the copper sulphides to the copper rougher concentrate and reduce the copper levels in the pyrite rougher concentrate. i?

16 SESSION A2 Optimization of Ore-dressing Processes and Energy Saving (continued)

I f Paper A2/1

THE IMPROVEMENT OF HYDROCYCLONING EFFICIENCY BY MODIFICATION OF THE VORTICITY FUNCTION OF THE FLOW FIELD

by

M.D. Brayshaw* and E.T. Woodburnf *Potchefstroom University for Christian Higher Education, South Africa fUniversity of Manchester, England

A numerical model using cylindrical polar coordinates (r, 0, z) was written for the rotational flow of a fluid in a cylindroconical hydrocydone based on the fluid-flow assumptions given below: (i) in viscidity, or viscous forces are considered insignificant, (ii) incompressibility, or the fluid density is unaffected by pressure, (iii) steadiness, or the solution is independent of time, (iv) axisymmetry, or the solution is independent of 0, and (v) homogeneity, or, effectively, the solids concentrations in the feed pulp are low, making the particle paths and streamlines coincident. For axisymmetry the azimuthal vorticity component, &>#, is given according to I equation (1):

du dw Tz ~ Tr- (1)

Two velocity components, u and w, written in terms of the Stokes stream function, , are given by equation (2) and identically satisfy continuity:

13* 1 3tf (2a,b) r az r or

When these expressions for u and w are inserted into equation (1) with otg/r defined as a unique function of \^, equation (3) is derived:

1 + ? a? (3)

17 Equation (3) was solved numerically by use of the Alternating Direction Implicit i technique. Fluid enters through a thin inlet slit circumscribing the full 360 degrees of * the circumference of the hydrocyclone so that the feed fluid is admitted in an axisymmetrical manner. It was found that the exact nature o(the/($) vorticity function effectively determined the resultant solution for 4'- A given functional form for^^) produced a solution for ^ that very closely approximated the spatial distribution of the three velocity components, u, v, and w, determined experimentally by Kelsall in 1952. Other f$) functions were incorporated in the model, and eacii was found to give its own unique solution for $. A method was developed for the computation of a theoretical Tromp or particle- separation-efficiency curve based on the concept of locus of zero vertical velocity and • on the other two velocity components implicit in the respective solution for ip. This method involved the equating of centrifugal and radial drag forces to give a particle at equilibrium on the locus. The theoretical efficiency curve consisted, essentially, in showing the change in size of this theoretical particle along the length of the locus. Solutions for \f> for different vorticity functions demonstrated further that this^^) function had a profound effect upon the theoretical efficiency curve. A gamma type °f/W function produced a much sharper efficiency curve than did the^^) characteristic of the flow of fluid in a conventional hydrocyclone. An analytical method was devised for the determination of the relationship between a given flow-field vorticity function and the velocity profile in the inlet to the hydrocyclone. This device laid the foundation for the testing in the laboratory of the findings that had been determined on a computer. Two cyclones were constructed. The first was a conventional single-entry hydrocyclone 3 inches in diameter. The second cyclone was the same as the first (i.e. diameter, length, outlets, etc.) apart from the inlet configuration. For this second hydrocyclone, seven inlets (the total cross-sectional area being exactly the same as for the single-inlet model) were spaced at 90-degree intervals round the circumference of the cylindrical section but at different axial heights. A given velocity profile in the inlet to this second hydrocyclone could be arranged by manipulation of the seven valves controlling the flow of fluid to each of the seven inlet pipes. The total flow of slurry, for a comparative run of the two cyclones, was the same in each hydrocyclone. A feed slurry was made from silica powder and water, and was mixed in a stirred tank situated beneath the two mounted cyclones. The conventional cyclone was then compared with the second, or multi-entry, hydrocyclone operating with a given predetermined inlet-velocity profile. Pulp densities of up to 40 per cent solids by mass were used, and in every instance the tailored velocity profile produced a much sharper 'cut' or separation-efficiency curve of the particles in the feed suspension. The laboratory work therefore verified the ideas brought out during the theoretical or development stages of this work, proving that a change in the vorticity component of the total flow field exerts some significant and beneficial effect upon the sharpness of particle separation in a hydrocyclone.

18 ., Paper A2/2 i i j CONTROL AND OPTIMIZATION OF THE OPERATION OF MILLING j CIRCUITS

i by

D.G. Hulbert and IJ. Barker Council for Mineral Technology South Africa Many of the problems encountered in the control and optimization of the operation of milling circuits are common to most of these circuits, yet there is still a wide diversity in the objectives and methods adopted. A number of milling-control projects have been undertaken by the Council for Mineral Technology (Mintek), and this paper is based I on the practical experience gained in these projects. The paper considers the fundamental - . aspects in the control of wet-milling circuits, with special emphasis on the area in which principles of both metallurgy and control engineering are required. The essence of 'good' milling in the metallurgical industry is analysed, and this is related to the requirements of a satisfactory control scheme. A milling circuit is ideally required to operate at a specified long-term rate of throughput, and give an optimum average product that results in optimum downstream processing. In most instances, i< measurements of the down-stream process or its efficiency cannot be incorporated into an automatic control scheme for the milling circuit, which is controlled as an independent unit. The optimization is then carried out by off-line adjustments of parameters or setpoints. Some commonly chosen control objectives may be inferior to others because they are ill-suited to the particular application. For example, the elimination of short-term fluctuations may result in less improvement than that achieved by the operation of the plant at slightly better average conditions. The importance of throughput and size distribution of the product is discussed in relation to the control scheme. The operation of a milling circuit at a fixed feedrate of solids is regarded as an unnecessary restriction of a potentially valuable control action, even when a particular long-term average throughput is required. Not all the important variables on a milling circuit can be measured direct or derived | from on-line measurements. Also, there are generally far fewer control actions available < j in practice than would be necessary for complete control of a muling circuit. The resulting : | compromises in a control scheme should therefore take advantage of the inherent > \ metallurgical properties of the circuit. For example, the lack of suitable measurements i of the underflow from a cyclone can be overcome to a reasonable extent by the use of measurements of the feed stream and a knowledge of the characteristics of the cyclone. The ultimate aim in the application of a control scheme is not just steady operation, but steady operation at the best operating conditions. The optimization of the operation

19 of a milling circuit after the application of a multivariable-control scheme is discussed. Particular reference is made to the project carried out at East Driefontein, where the size of the product is controlled and two other setpoints are selected according to a practical scheme that results in the maximum throughput of solids at the required size. The derivation of the optimization scheme is described. Part of the optimization also included minor modifications to the milling circuit where physical limits were encountered and easily identified because of the well-controlled operation of the plant. Significant benefits can be obtained by the well-planned application of control and optimization techniques.

Paper A2/3

THE MEASUREMENT OF PARAMETERS DESCRIBING THE DYNAMIC BEHAVIOUR OF THE LOAD IN A GRINDING MILL

by

M.H. Moys University of the Witwatersrand South Africa

A 2,5 kW pilot grinding installation is described. The mill has a variable speed (10 to 100 per cent of critical speed), and all variables such as flowrate, densities, torque, and mass of load can be measured or controlled with a high degree of accuracy. (Errors are less than 1 per cent of the value of the variable in most cases.) A method for measuring the dynamic angle of repose has been developed. The conductivity between the mill shell and a probe inserted through the mill shell is measured and communicated to the outside world via slip rings. A switch operated by a cam on the slip-ring assembly is used to indicate the time when the conductivity probe passes a certain point in its circle of rotation. These signals are analysed using a process-control computer to provide estimates of the speed of rotation, N(r/min), of the mill, the period of immersion of the probe in the load, r^, and the dynamic angle of repose of the load, o. The mill was operated as a ball mill with a grinding medium of 60 to 350 kg of steel. The above variables were measured under various operating conditions. With 300 kg of medium, 0 increases steadily from 31° at 10 per cent of critical speed (JVC) to 60° at 80 per cent Nc. At speeds above 75 to 80 per cent Nc, cataracting of the load occurs and it becomes difficult to define or measure 0. Increasing the volume of the load from 30 to 35 per cent of mill volume had a negligible effect on 0. As the volume of slurry is increased from 0 to 110 per cent of the volume of the voids in the medium, a downward trend in 0 is observed: at 77 per cent Nc, 0 decreases by approximately 7°. With a load volume of less than 30 per cent of mill volume, the torque developed by the mill does not vary by more than 3 per cent as the speed varies from 10 to 85

20 per cent Nc. However, using a simple model which ignores dynamic energy recovery from the load to the mill shell, it can be shown that torque should be proportional to sin 0. Since 6 varies over a wide range as mentioned above, it is concluded that energy recovery takes place at a rate sufficient to maintain constant torque over the above speed range. For speeds greater than 85 per cent Nc, torque decreases rapidly as energy recovery from the cataracting load becomes increasingly important. Load volume is strongly correlated with TL at a given N, and TL can thus be used to provide a dynamic estimate of the volume of the load. The measurement of 6 and r^on industrial-scale mills is discussed. The signal will be transmitted radiometrically rather than via slip rings; since only on/off signals need be transmitted, this should not present serious problems. The conductivity probe will be mounted in a liner bolt and can be renewed during regular maintenance downtime. The availability of a measurement of 0 and r^ will provide useful information for dynamic mill control, particularly for the control of autogenous, semi-autogenous, and pebble mills.

Paper A2/4

THE EFFECTS OF CHANGES IN LINER DESIGN AND MILL SPEED ON ROD-MILL PARAMETERS

by

D.D. Howat and L.A. Vermeulen Council for Mineral Technology South Africa The use of rod mills for primary grinding is a comparatively recent development in the gold mines of South Africa, the first two rod mills (each 2,44 m by 3,66 m) having been installed at a mine in the Orange Free State in 1951. Since then the merits of rod mills for the primary grinding of gold ores have become widely known and recognized. The feed, normally a screened product from the secondary crushers, is nominally up to 19mm. Commonly, the mills are lined with backing blocks recessed to permit lifter bars to be bolted through the blocks to the shell. With a relatively coarse feed and the heavy pounding produced by steel rods of 100 mm diameter, the consumption of lining metal is relatively high. The other main cost item in milling is electric power, which, with the rapid rise in price in recent years, is now a dominant factor. Tests were carried out at three mines on the effects of changes in the number of lines, heights, and spacing of lifter bars in attempts (a) to determine the most cost-effective wear-resistant alloys, and (b) to reduce the consumption of liner metal and of electric power per ton milled. The investigation ranged from the testing of relatively smooth liners to the doubling

21 of the number of lines of lifter bars and to the use of alternating lines of higher and lower lifter bars. Of the materials made available for the tests, austenitic manganese-steel castings offered decided advantages, particularly under local conditions. An increase in the number of lines of lifter bars and in the height of the bars yielded maaked savings in the consumption of liner metal and electric power, and permitted a larger throughput. The relative efficiencies in the utilization of electric power in mills operating under different conditions was calculated, sometimes with surprising results. By the use of a laboratory mill, the magnitude of slip of rods in a mill was studied, as were the effects of changes in operating speed on the trajectories of the rods.

\

22 SESSION A3 Optimization of Ore-dressing Processes and Energy Saving (continued)

Paper A3/1

FLOTATION MACHINE SCALE-UP: THE PATH TOWARDS j IMPROVED PROCESS ECONOMICS i by

V.R. Degner WEMCO U.S.A.

The trend in the minerals industry towards very large flotation machines (exceeding 10 m3) began in the late 1960s and continues today at an ever-increasing pace. The impetus for large flotation machines is found in the economical utilization of floor space, power requirements, and operational/maintenance manpower needs of the concentrator. For example, a 50 per cent increase in the dry tonnage processed per unit of floor area follows the replacement of 14 m3 machines with an equivalent flotation capacity of 28 m3 machines. The utilization of operational manpower is similarly enhanced when a high-tonnage concentrator is equipped with a fewer number of larger machines. In general, as the process-design tonnage increases, the benefits of the larger flotation machine become more evident. Consistent with current plant design and operation practice, individual flotation cells exceeding 28 m3 (1000 ft3) are rapidly becoming more familiar. WEMCO's development efforts in large flotation machines progressed from the 8,5 m3 machine, introduced to the market in 1967, through the 14 m3 (1971) and 28m3 (1978), to the 43 m3 machine, which began in-plant testing in 1982. In this paper, the analytical methods used to hydrodynamically scale up the technology of flotation-machine design are reviewed in detail. In addition, the analysis is extended to the configuration of the next larger flotation-machine design. The configuration of any WEMCO large flotation machine is aimed at mutually satisfying the process requirements for (1) good air-pulp contact and mixing, (2) stable froth surface, (3) adequate solids suspension, and (4) acceptable froth removal from cell to launder. This paper identifies, in detail, the five hydrodynamic scale-up parameters which govern the configuration of large flotation machines. Four proven (existing) designs for machines ranging in size from 8,5 to 42 m3 are reviewed, and the design of a significantly larger machine is identified. The five hydrodynamic design parameters which serve to define the configuration of flotation machines are (a) velocity of air escaping into the froth surface, (b) specific circulation or pulp circulation/feed rate,

23 (c) air-pulp mixing (residence time of air-pulp contact), (d) mechanism power number, and (e) (absolute) pulp circulation at the cell floor. The importance of the velocity of escaping air, which affects the surface turbulence of both self-aerating and external-air types of flotation machine, is discussed for a range of cell sizes. The relationship between cell configuration (size) and mechanism power input, as identified by a threshold for air 'geysering' at the froth surface, is a dominant consideration in the design and development of flotation machines. Evidence in support of this criterion is summarized. Pulp 'circulation intensity' (by definition, internal circulation through the mechanism cell volume) is equivalent to the number of theoretical 'turns' the pulp will take through the mechanism during its residence in a given flotation cell. The larger the circulation intensity, the higher the probability of contact between the air bubbles and floatable species and of subsequent recovery of the species. Air-pulp 'mixing' is quantified by the mechanism power number and the air-pulp residence time within the mixing zone, and suggests the magnitude of contact probability between air bubble and floatable species. A review is given of analytical means by which the existing designs of flotation-machine product lines are utilized to provide a statistical base for rational extrapolation to larger equipment, in the light of the five specific criteria identified above. In addition, the results of current research-and-development studies, aimed at relating the metallurgical performance to key hydrodynamic considerations, is summarized. A key to establishing the hydrodynamic scale-up basis for a given class of flotation machine is the quantitative relationship between air distribution or air-transfer capability, liquid (pulp) circulation, and power for a given geometry of flotation mechanism and operating condition (i.e. rotor speed). These relationships have been developed over several years, and have now been organized into a set of empirical equations covering all the machines included in the current product line. This data base becomes the rational means by which the configuration of the flotation mechanism for any future new machine can be rapidly established with confidence. Specific plant installations and established metallurgical performance are summarized briefly to show the economic benefits of the large flotation machine in the contemporary concentrator application. In summary, the spatial and power requirements (i.e. design) of the large flotation machines currently available and planned for the future are related, on the basis of rational hydrodynamic design scale-up criteria, to illustrate potential improvements in system economics. Recent studies of flotation kinetics by WEMCO are used to quantify metallurgical performance in complementary evaluations of process scale-up.

24 Paper A3/2 THE INFLUENCE OF POLYMERIC DEPRESSANTS ON THE ADSORPTION OF THIOL COLLECTORS IN SULPHIDE FLOTATION

by

E. Steenberg and P.J. Harris Council for Mineral Technology South Africa Polymeric depressants are used in the selective flotation of valuable sulphides, sometimes containing gold or platinum, from talcose gangue. However, no information is available — . I about the adsorption of these reagents on the sulphide minerals themselves, or about I the influence of these polymers on the adsorption of thiol collectors. This paper describes the results obtained from a study of these topics. Use was made of two thiol collectors—sodium ethyl xanthate (at a pH value of 7,5) and sodium mercaptobenzothiazole (at a pH value of 4,0)—and three polysaccharide depressants—sodium carboxymethylcellulose (CMC) (molecular mass 9 x 106), a modified guar gum (molecular mass 1,1 x 10s), and a potato starch (molecular mass 1,9 x 107)—and their adsorption behaviour and influence on the floatability of pure galena and pure pyrite were examined. Initial tests showed that no adsorption of the thiol collectors could be detected on pure talc (chosen as a typical gangue mineral), whereas the polymeric depressants were adsorbed on the two sulphide minerals. Adsorption isotherms of the three depressants on galena and pyrite were determined at pH values of 7,5 and 4,0, and the influence t of the polymers on the zeta potentials of the two minerals was measured. The nature of the three polymers is very different: CMC is strongly negatively charged, the guar gum has been modified to include some anionic substituted groups, and the starch has a very low degree of substitution. However, they are all adsorbed to a significant extent on galena at neutral and acid pH values. There are indications that the presence of lead ions in solution affect the adsorption of both guar gum and starch. At pH 7,5 the zeta-potential measurements show that the positively charged lead hydroxy species is included in the adsorbed polymer layer. The adsorption of the polymers on pyrite was more pH-dependent than that observed with galena. At the higher pH values, the adsorption of strongly dispersing polymers (e.g. CMC) on the dispersed ferric hydroxide (formed by the oxidation of pyrite) interfered with the adsorption of the polymer on the pyrite itself. For both minerals, the zeta potentials were reduced in the presence of guar gum and starch, and increased in a negative direction in the presence of CMC. In the second phase of the investigation, the polymers were adsorbed before the collector was added, and the adsorption of the depressant and the collector were monitored for a further period. The order of reagent addition was then reversed, and the depressant was adsorbed on a sulphide surface previously exposed to collector. The adsorption of xanthate on galena is inhibited by the presence of the previously adsorbed guar gum or starch. Indications are that the polymers interact strongly with the ionic lead species present, thus reducing the formation of lead xanthate. On the

25 addition of xanthate to a galena surface previously exposed to CMC, the polymer is apparently completely desorbed and the adsorption of xanthate can proceed unhindered. j If xanthate is added before the addition of the polymer, adsorption of the polymer can { , still occur. In comparison with the results obtained with the reverse sequence of addition, 1 the adsorption of guar gum is not inhibited, the adsorption of starch is inhibited to \ a certain extent, and no adsorption of CMC appears to take place. j The presence of previously adsorbed guar gum and starch does not interfere with : '•• the adsorption of xanthate on pyrite. As found with galena, the presence of adsorbed xanthate on pyrite does not inhibit the adsorption of guar gum but does reduce the amount of starch adsorbed. Owing to the adsorption of CMC on the dispersed ferric hydroxide, the effect of CMC on the adsorption of xanthate on pyrite could not be established. At pH 4,0 the reaction of mercaptobenzothiazole with pyrite is not affected by prior adsorption of guar gum, but it is inhibited by the presence of CMC. A reduced amount of guar gum and CMC can be adsorbed on a pyrite surface previously exposed to mercaptobenzothiazole. The results of flotation tests conducted in a single-bubble microflotation cell indicate that the mineral particles are floatable even if a certain amount of polymer has been | adsorbed on the surface. However, if sufficient polymer has been adsorbed, the particles '_ can be depressed even if collector has been adsorbed. Thus, the floatability depends ,: on the ratio of collector to depressant adsorbed at the surface. The trends observed ! in these adsorption studies are parallel to those found in the flotation tests.

Paper A3/3 THE HYDROPHOBIC CHARACTER OF SEMI-SOLUBLE SALT MINERALS WITH OLEATE AS COLLECTOR

by

J.D. Miller and M. Misra University of Utah U.S.A. One of the most important collectors in the flotation of semi-soluble salt minerals is oleic acid. Although considerable research has been reported for most of these minerals on the nature of the oleate adsorption reaction (chemisorption/chemical reaction) and the flotation response, little attention has been given to the hydrophobic character of these systems. In this regard, recent measurements of contact angles and induction times for selected systems involving oleate and semi-soluble salt minerals indicate that ! important surface reactions can change the hydrophobicity of the system. These reactions were studied by spectroscopic techniques and were found to be dependent on the ' composition of the solution and gas phases and on the temperature. In the case of the fluorite-oleate system, the hydrophobicity, as assessed by both equilibrium and kinetic measurements, was found to be significantly dependent on

26 temperature, oxygen partial pressure, and oleate concentration. Table 1 compares the effect of these variables on the hydrophobicity of fluorite in the presence of oleate. Such a dependence is not observed for saturated fatty acid collectors.

TABLE 1

Measurements of air-bubble attachment at the fluorite surface in the presence of oleate

Conditions: Conditioning time 10 mm, pH value 8,1 Estimated Oleate attachment Gas concentration Temperature Contact angle time phase M °C degree s Air io-5 22 68 to 70 0,5 Air io-5 60 90 to 95 0,002 to 0,004 Oxygen 10"5 22 90 to 92 0,002 to 0,004 Air io-4 22 85 to 87 0,004 to 0,005 Air 10"3 22 90 to 92 0,002 to 0,004 (bubble leaves surface after 0,2 s)

The effects of temperature, partial pressure of oxygen, and other solution-chemistry conditions alter the hydrophobic state of the fluorite-oleate surface. These results, together with the results for other semi-soluble salt minerals and the reports of previous investigations, suggest that the interaction of adsorbed unsaturated fatty acid collectors is important. This interaction seems to be a rather complex phenomenon and may involve oxidation with subsequent polymerization of the unsaturated radical, resulting in improved hydrophobicity and increased flotation rate. A tentative mechanism for such an interaction of unsaturated fatty acids can be proposed in the light of the research done in the fat and oil industries, which may be, in a restricted sense, applicable to systems involving non-sulphide minerals. Recent flotation research on the magnesite system at the University of British Columbia gives evidence that these reactions do occur. In this regard, a possible reaction sequence would be hydroperoxide, epoxide, and oxirane polymerization. Finally, in addition to high-speed video recordings of bubble attachment, other video recordings clearly show that, for excessive additions of oleate, the buoyant force of the air bubble can tear the film of calcium oleate from the surface. These observations help to explain the poor flotation response that has been observed by some investigators for I f certain semi-soluble salt minerals at high levels of oleate additions.

27 Paper A3/4 1 i A STUDY OF THE SELECTIVE FLOTATION OF FLUORITE FROM CALCITE ) BY THE USE OF A SINGLE BUBBLE-STREAM MICROFLOTATION CELL 1 by

E.W. Giesekke and P.J. Harris Council for Mineral Technology South Africa

Adsorption and microflotation studies were conducted on fluorite and calcite so that an explanation could be obtained of how a selective recovery of fluorite over calcite is achieved during flotation with fatty acid. Sodium oleate was used as a collector and wattlebark extract (WBE) as a depressant. The influence of Ca2+ ions (apart from the Ca2 + ions originating from the dissolution of the minerals) on the adsorption of the various reagents was also studied. The microflotation studies were carried out in a single bubble-stream cell in which the gas flowrate was carefully controlled. The bubble-size distributions, and consequently the total surface area of gas available for flotation, were measured in each test. Under these carefully controlled conditions, the flotation behaviour of the two minerals with varying additions of reagents could be accurately compared. In the absence of additional Ca2+ ions, the added sodium oleate collector forms calcium oleate, CaOl2, on both minerals in multilayer quantities, to a greater extent on fluorite than on calcite. These multilayers build up owing to the continuous release from the mineral surface of Ca2+ ions, which interact with the sodium oleate. In addition, significant amounts of CaOl2 are precipitated in solution, and the concentration of free Ca2+ ions in solution is maintained at approximately 3 p.p.m. The nature of CaOl2 on fluorite is different from that on calcite, and forms the basis for the selective recovery of fluorite. WBE improves the selectivity by decreasing the amount of CaOl2 adsorbed on calcite to a relatively greater extent iCian on fluorite. If additional Ca2+ ions are present in solution before the oleate is added (e.g. with hard water), the bulk of the CaOl2 is formed in solution and not at the mineral surface. At room temperature under these circumstances, the amount of CaOl2 attached to both mineral surfaces is very similar and much less than previously mentioned, and consequently no selectivity is observed. Both minerals are easily depressed by WBE. An increase in the pulp temperature to about 70 °C in the presence of WBE enhances the adsorption of CaOI2 on fluorite, leading to an improved recovery. These observations indicate that the CaOl2 adsorbed on calcite is different from that adsorbed on fluorite, and that the amount of GaOl2 adsorbed on either mineral is not an indication of its floatability. On calcite the results suggest that the CaOl2 is essentially a loosely held precipitate, whereas on fluorite a greater proportion of the CaOl2 is more strongly bound, particularly at higher temperatures; and that this strongly bound form can be present in greater than monolayer coverage.

28 SESSION A4 Treatment of Fines, Tailings, and Low-grade Ores • ! Paper A4/1 ! h { . SURFACE INTERACTIONS IN FINE PARTICLE FLOTATION i ! ] T by P. Somasundaran Columbia University U.S.A. The problems involved in the processing of fines are enormous in that as much as 20 ••' to 50 per cent of the mineral values is lost during the processing of many ores. Several factors such as decrease in mass and increase in surface area have been considered in the past to be responsible for this. In this paper, the role of alterations in morphology, : mineralogy, and surface chemical composition in the fine and ultrafine size range will be considered. i Morphological differences can result from differences in the fracture of fines by virtue of their grain-size distribution or crystal structure, differences in grinding mechanisms between coarse and fine grinding, and differences in mineralogical composition as a function of particle size. Evidence presented in this paper indicates that the morphology of fines can, indeed, be different from that of coarse particles. For example, increased angularity in the slimes range was observed for the washings of coarse hematite particles. Fines of quartz prepared by abrasion appeared to have undergone more erosion in comparison with quartz particles from an impact tester. Prolonged grinding of minerals can produce amorphousness, polymorphic transitions, and even solid-state reactions. Differences in morphology can have a significant effect on flotation. There is sufficient evidence in the literature to show that the wettability of particles is strongly dependent on roughness. Measurements of contact angles have shown that a totally hydrophobic : material can be made hydrophilic by being roughened. The second major characteristic of fines that gives rise to problems is mineralogical alteration in this size range. The effect of such mineralogical alteration is best seen with phosphates. While coarse phosphate is mostly apatite, phosphate in slimes is present, • not only in the form of apatite, but also in the form of wavellite, which will respond j poorly to separation by flotation. \ A more serious problem arises from possible changes in surface-chemical composition due to oxidation of the surface, interactions in the surface region leading to precipitation, i or to coating of the surface by precipitates or slimes. For example, fines of sulphides ; are found to be more oxidized than coarse sulphides. With non-sulphides, surface- ! chemical alterations can often result from precipitation of inorganic, organic, and even polymeric compounds. Interaction between such surface layers more than anything else controls the flotation and flocculation behaviour of fines. Results obtained for the calcite- • apatite system clearly illustrate die effect of such surface-chemical alterations. Conditions predicted for the separation of calcite from apatite on the basis of single-mineral tests •

29 rarely work with mixed-mineral systems. Zeta-potential results of apatite fines in water j and in calcite supernatant show that the point of zero charge of apatite in calcite ! supernatant shifts to that of calcite, suggesting that the surface of apatite has been | j converted to that of calcite. Under such conditions, a flotation separation based on the i : surface properties of apatites is not likely to work. In fact, the flotation of apatite by | the use of oleate is affected markedly when calcite supernatant is used instead of water. : When collector adsorbs or precipitates on the surface of minerals, there are surface forces that are primarily responsible for the separation processes, particularly carrier flotation and flotation processes using polymers. Models for interactions between particles and particles, and between particles and bubbles, based only on electrostatic and dispersive forces, cannot explain the behaviour of these systems. In carrier flotation, where anatase impurity is floated from clay by the use of coarse calcite as carrier particles and oleate as collector, the mechanism responsible for the process is enhanced aggregation between anatase and calcite and almost no aggregation between clay and calcite. A detailed analysis of the system showed the interaction between adsorbed oleate layers to be responsible for the aggregation in this system. Selective aggregation of anatase with calcite is attributed to weaker attachment of oleate to the clay surface compared j with the anatase surface such that the clay-calcite aggregates formed are not stable at the high agitation intensities used in carrier flotation. In modified flotation processes such as floc-flotation, there are many interactions between surface layers that are of extreme importance. Polymers can affect flotation by depressing it when they are added to flotation systems. Normally, a chemical that depresses flotation is considered to act by preventing the adsorption of collectors. A study of the quartz-dodecylamine-cationic polyacrylamide system showed that the I polymer depressed flotation without any effect on the adsorption of the surfactant itself. Zeta-potential results showed that, whenever the polymer adsorbs, irrespective of whether the amine adsorbs, it is the zeta potential of the polymer that is always recorded. The model that accounts for this depression is such that adsorbed amine is masked by massive polymer species leaving the surface of the particle essentially cationic and thereby hydrophilic. This model predicts the flotation of quartz using anionic surfactant in the presence of a cationic polymer, even though normally such a surfactant would not float quartz. This was indeed observed for the quartz-dodecylsulphonate-cationic polyacrylamide system. Evidently, therefore, there are many interactions between adsorbed layers, governed by their properties such as strength of binding and configuration, which have to be taken into account for a full understanding of the operations in the processing of fine particles.

30 Paper A4/2

PYRITE FLOTATION WITH DIETHYLDITHIOPHOSPHATE AND MERCAPTOBENZTHIAZOLE: AN ELECTROCHEMICAL STUDY by

D.R. Groot Council for Mineral Technology South Africa The flotation of pyrite with thiol collectors at acid pH levels is important in the processing of tailings and leach residues. Although potassium diethyldithiophosphate (DTP) and sodium mercaptobenzthiazole (MBT) are often used as collectors, the mechanism of their interaction with pyrite is not well known, and optimization of the flotation is often difficult. Microflotation studies, coupled with electrochemical measurements, were undertaken to provide a better insight into the reactions that occur. Experiments under various conditions, with potential monitoring of several pyrite electrodes, showed that the ultimate flotation recovery with DTP as collector is related linearly to the electrode potentials. However, when MBT was used as the collector, no trends were observed. It was thus established that the flotation recovery depends on the percentage of mineral particles that have the appropriate surface potential for attaining floatability. Measurement of the potentials of individual pyrite particles obtained by the crushing of a single lump of ore showed a range of values under a given set of conditions, with the mean potential and the range depending, for instance, on oxygen concentration and the presence and type of collector used. Cydovoltammetric investigations, coupled with in situ measurements of contact angles, on pyrite and noble-metal electrodes in DTP solutions showed that the degree of hydrophobicity of the electrode surface is potential-dependent. If the electrode is at a sufficiently positive potential, reversible adsorption of the collector occurs. At far higher positive potentials, oxidation of DTP to the dithiolate occurs on the surface, resulting in a greater degree of hydrophobicity than when only adsorbed collector ions are present. These results thus provide an explanation for the linear trend observed in flotation recoveries versus monitoring electrode potential. The above investigations all provided indirect information about the identities of the surface species. Attenuated total-reflectance infrared spectroscopy on organic solvent extracts from pyrite that had been floated with DTP and MBT showed spectra that correlated well with the spectra obtained from DTP and MBT that had been oxidized chemically. This method could not, however, be used to give direct information about surface species other than these oxidized collectors. Through microflotation and cyclovoltammetric studies, the depressant effects of cyanide and sulphur dioxide were investigated. The action of cyanide could be explained by its strong adsorption, resulting in a decreased interaction between the collector and oxygen with the mineral. However, sulphur dioxide does not adsorb strongly, but it acts as a reducing agent and thus depresses the recovery. The insights gained so far through this electrochemical approach to flotation problems are proving the value of this type of study.

31 Paper A4/3 IMPROVEMENTS IN THE FLOTATION OF LOW-GRADE FLUORSPAR ORES

by

P.G. Kilian Gencor Group Laboratories (East) South Africa The paper describes the development of a modified flotation process using fatty acid collector for the recovery of fluorspar from iow-grade ores. The difference between the conventional process and that described is essentially in the conditioning stage. Most of the work, including all the continuous testing, was carried out on low-grade, brecciated fluorspar from the Zeerust area, in which dolomite is the predominant gangue mineral. Some confirmatory work is continuing on a fluorspar ore in which silica is the predominant gangue constituent. The objective was to optimize a flotation process to obtain the maximum recovery at acid grade (over 97 per cent CaF2). Preliminary bench flotation tests were conducted according to the conventional method, with variations of temperature, collector addition, pH modifier, particle size, and depressant. At least six stages of cleaning were required to reach acid grade at optimum conditions. The best recovery in a final concentrate of acid grade was iess than 80 per cent. The process was found to be extremely dependent on the addition of collector and depressant, and on the quality of these reagents. A small pilot plant was designed and built to test the process on a continuous basis. All the attempts were unsuccessful owing to inadequate control of the sensitive parameter:;. Following up on work of a more theoretical nature on the bonding of fatty acid to fluorspar, the work was continued on a bench scale using the modified method of conditioning the pulp at boiling point. In this reaction, the fatty acid collector attaches itself to the surface of the fluorspar particles in a semi-permanent bond. Once this bond has been achieved, the hydrophobic properties can be removed only by vigorous scrubbing with acid or organic solvent. The process has the following advantages. (1) Collector-to-mineral bonding takes place rapidly at boiling point (96 °C). The reaction is complete within 5 minutes. (2) The transformation is visual. (3) The final concentrate grade is less dependent on collector or depressant addition. (4) The recovery of fluorspar is almost total in the rougher stage. (5) Fewer cleaning stages are necessary because of the increased selectivity. (6) The process is not sensitive to operator differences. (7) Owing to a lower water content in the conditioning stage, the energy requirements for heating the pulp are lower than in the conventional process. (The calculations involving the energy input are presented.) At optimum, the flotation yielded a recovery of 93 per cent CaF2 in a final acid-

32 grade concentrate from a feed grade of 12 per cent. The efficiency of the process resulted in comparatively low concentrations of calcium fluoride in the rougher tailings, making analysis difficult. The analytical methods employed arc briefly discussed. Several attempts were made on a continuous pilot plant to achieve the results of the bench tests. The limited success was due to several problems, some of which involved (a) obtaining a reasonably steady feed of pulp thickened to 70 per cent solids, (b) short-circuiting of the pulp in the single-stage conditioner, and (c) inability to remove the extremely dense froth from the flotation cells. A description of the methods used to overcome these problems is given. Two final runs on the continuous plant with recycling of both solids and water were carried out. The objective of a 90 per cent recovery at acid grade was achieved without difficulty. Bench tests on fluorspar ores in a different gangue matrix have to date shown promising results. Recoveries in excess of 90 per cent are possible on many ore types.

Paper A4/4 GOLD RECOVERY BY FLOTATION AT BOUGAINVILLE COPPER LTD

by

N.C. CSarke and E.W. Beermann Bougainville Copper Ltd Papua New Guinea

Bougainville Copper Ltd (BCL) currently produces about 18 000 kg of gold per year, which is recovered by flotation into a copper concentrate. The recovery of copper is 83,5 per cent and of gold, 72 per cent. Losses of gold to the tailings are 7000kg per year at a tailings grade of 0,15 g/t Au. When the feasibility study for the Panguna mine was completed in 1969, gold had a fixed price of 38 U.S. dollars per ounce. It was expected to contribute 10 to 14 per cent of gross sales revenue — important but not vital to mine profitability. Mine planning and metallurgical development after start-up in 1972 was geared to predicting and improving copper production. A high degree of predictability and control has been achieved as a result, for copper. The rising price of gold through the 1970s compared with copper has resulted in gold now contributing very close to 50 per cent of BCL's revenue. The changes made to the planning procedures to fully incorporate gold are discussed. However, gold recovery cannot be controlled in the plant, and little is known of the response of gold to flotation process variables. Nor has the reason for the comparatively low gold recovery compared with that of.copper been understood. To provide information on these factors and to identify means of improving gold recovery, an investigation was conducted into the modes of occurrence of gold in the feed and the recovery of gold by size and by type. An initial belief, arising from analyses of hand specimens and electron-microprobe scans of flotation concentrates, was that the majority of the gold was in solid solution

33 in copper sulphides, and that the liberated gold not recovered in flotation was the major source of gold losses. A corduroy strake treating part of the tailings stream recovered up to 14 per cent of the gold in the tailings, and a pilot plant was built to experiment with more suitable means of gravity concentration. * Gold recove-ies from portions of the main tailings stream averaged 5 per cent, with concentrate grades of 0,3 g/t. The equipment used was shown to be operating satisfactorily, and the low recovery was due to the lack of liberated gold in the tailings in a size range amenable to conventional gravity-recovery techniques. It was also shown, from shaking table and hydrocydone tests, that there was a high proportion of liberated gold in the final concentrate, and thus that liberated gold was recovered by flotation. An extensive study of the mill products using the optical microscope showed that gold occurred in or attached to gangue materials, liberated, or enclosed in or attached to copper sulphides. Later work found gold also enclosed in or attached to pyrite. Optical microscopy cannot give quantitative results on gold occurrence in low-grade gold ores. The Australian Mineral Development Laboratories were commissioned to examine scavenger flotation tailings. This work was later extended to include cleaner circuit tailings, which is the major pyrite-bearing stream. The approach used is described. The intrinsic gold levels of the sulphide minerals varied with size, and probably with ore type. Chalcopyrite coarser than 7 pm contained between 1 and 16 p.p.m. of gold. Pyrite coarser than about 8/un contained 1 to 4p.p.m. of gold. The gold content of bornite is uncertain but is probably about 40 p.p.m. Below the sizes given, the gold content reduces sharply. Chalcopyrite comprises 1,2 percent of the flotation feed, pyrite \ 0,6 per cent, and bornite 0,1 per cent. In scavenger tailings in the sample examined, 39 per cent of the gold was locked with silicates, 4 per cent was contained in sulphides, and 55 per cent was believed to be liberated. The information on the mode of gold occurrence by size and by type was obtained on samples of flotation feed and scavenger tailing taken simultaneously. The data, obtained in 1980, are given in Table 1. TABLE 1

Gold recovery in flotation by type

Distn, % Recovery to concentrate, % Au in Au in Au, Size Mass Au sul- sili- liber- pm % Cu Au Cu total phides cates ated >75 42 30 31 74 67 74 51 84 <75 58 70 69 96 83 96 83 79 Total 100 100 100 91 78 91 61 80

The present gold recoveries, at 72 per cent, are lower than shown in Table 1, and the change is caused by an increase in the proportion of gold locked in silicates. The

34 good flotation recovery of liberated gold above 75 /*m in size makes it unlikely that gravity-recovery methods will substantially improve the gold recovery. The reason for the apparent decline in the recovery of liberated gold below 75 jtm is uncertain. Current research work is described that aims to elucidate this question and identify the options available for improving recovery.

f

35 SESSION A5 Treatment of Fines, Tailings, and Low-grade Ores (continued)

Ji Paper A5/1

THE ROLE OF WASTE SORTING IN THE SOUTH AFRICAN GOLD MINING INDUSTRY

by

J.S. Freer and R.C. Bohme General Mining Union Corporation Ltd South Africa Waste sorting is a process whereby gangue is separated from ore on a piece-by-piece basis, leaving an upgraded fraction for further treatment. It has the advantage of application near the head of the extraction process, before any significant costs are incurred for comminution. It might even be applied close to the mining face to reduce the cost of transport, or to increase shaft capacity, while the waste could be stowed for stope support. Its potential for use in the South African gold-mining industry is enhanced by two factors: (1) the deposition of gold, uranium, and pyrite values in reefs which are often narrower than the minimum width required for stoping, so that an additional thickness of country rock above and/or below the reef band has to be mined with it as unwanted dilution; (2) the heterogeneous distribution of these values within the reef so that, on blasting, large numbers of waste-rock pieces are generated as internal waste. The real potential for sorting is only now becoming clearly understood. It can best be established by determining the mass number of pieces of gangue in the size range big enough to sort. This is done by a process of weighing and assaying (for gold and U3O8) every rock piece making up the population of bulk samples of sized ore taken from large composites of run-of-mine ore, and analysing the results by computer. The most striking conclusion is the high percentage mass of gangue (65 to 75 per cent), even in those instances where the channel width exceeds the minimum stoping width, indicating the large amount of internal waste liberated by blasting. In cases where the channel width is narrower than the stoping width, the potential for waste rejection probably often exceeds 80 per cent. As sorting can be applied economically only to size ranges above 25 mm, waste rejection can be around 40 per cent and even as high as 50 per cent by mass of run-of-mine ore. Having established the waste:ore ratio at any given cut-off point for gold and/or uranium in this way, what parameters can be used for achieving the potential? The appearance of the rocks, on which both manual and photometric sorting rely, does not approach the potential, but it is more closely achieved by radiometric sorting. This has been used effectively for sorting uranium ores, and is now applied equally successfully to gold ores in some of those instances where the association of uranium with gold on a particle-by-particle basis is adequate for the purpose.

36 However, the greater proportion of the potential for sorting is from those ores in which the association of uranium with gold is not sufficiently faithful to provide an acceptable criterion for the recovery of the gold. For this purpose, natural radiometry is inadequate. To meet this challenge, General Mining Union Corporation Ltd has sponsored a research programme aimed at neutron activation of the gold contained in the rock pieces and then sorting based on the gamma activity thereby induced. A suitable nuclear reaction has been identified in which sized ore is irradiated by fast neutrons produced by a neutron generator. The irradiated rock pieces are moved rapidly tc an array of scintillation counters, which are monitored by computer to derive the gold grade of each rock piece and to actuate an accept-reject mechanism according to the same principle as the radiometric sorter. A comparison of the approximate costs of the following flowsheet options is attempted: (1) run-of-mine milling, leaching, and recovery from the whole ore; (2) screening (or screening and sorting), followed by stage comminution, leaching, and recovery from the accept fraction with or without treatment of the reject fraction; (3) screening (or screening and sorting), followed by autogenous milling of the accept fraction, leaching, and recovery from die accept fraction with or without treatment of the reject fraction; (4) run-of-mine milling in closed circuit with the sorting of certain size fractions from the mill load, leaching, and recovery from the milled product; (5) sorting underground with the objective of using rejected waste for stope support. The last option is compared with underground milling, flotation of a bulk concentrate for pumping to surface, and stowing of the flotation tailings as stope support.

Paper A5/2

SELECTIVE MAGNETIC FLOCCULATION OF FINE, WEAKLY MAGNETIC MINERALS

by

J. Svoboda Council for Mineral Technology South Africa

The processing of fine-grained ores frequently results in the production of ultrafine particles that respond poorly to conventional separation techniques. The formation of large units by selective tlocculation is a promising way to overcome the limitations of physical separation techniques. A suspension of solid particles can be destabilized, and floes can be formed either by chemical agents or by external forces. Flocculation by the external magnetic field offers a wide spectrum of possibilities of inducing selective agglomeration even of particles that are very weakly magnetic.

37 The stability conditions of a suspension of mineral particles can be determined from the shape of the curve for the total interaction energy of the particles. Under certain assumptions, the total interaction energy of small identical weakly magnetic particles (paramagnetic or diamagnetic) placed in the external magnetic field can be written as

exp[-T(5-2)] 2 s* -4 32II2 X2 In ? where 6 is the dielectric constant of the medium, a and x are the particle radius and magnetic susceptibility respectively, &, is the potential at the surface of a particle, A is the Hamaker constant, B is the magnetic induction, p0 is the permeability of vacuum, s is the normalized interparticle distance s = r/a, and r is the interparticle distance r = K'a, where K is the Debye-Hiickel reciprocal-length parameter. A curve of total interaction energy, VT, as a function of interparticle distance is shown in Figure 1. Particles can flocculate into the primary minimum provided the potential barrier Vmn is sufficiently low. Such a barrier can be suppressed by the placing of a system of particles into a sufficiently strong magnetic field in which the magnitude of the field depends on the radius, susceptibility, and surface properties of the particles.

FIGURE 1. The general form of the total potential energy for a pair of interacting colloidal particles

On the other hand, particles can also flocculate into the secondary minimum, Vmin, if it exists. In this instance, a potential barrier does not need to be reduced, and a sufficient condition for flocculation into the secondary minimum is that the potential well should be sufficiently deep. A threshold magnetic field at which the potential barrier separating two interacting particles disappears can be determined from the condition that the stability ratio for the primary minimum, Wp, should be smaller or equal to unity:

38 \Vr]ds jexp [Tf\7 Similarly, there is a sufficiently deep secondary minimum when the stability ratio for the secondary minimum, W,, is smaller than or equal to unity: J [itf]? W. - •*[w] The ratios of threshold magnetic fields for rapid flocculation into primary (2J£) and secondary (5f) minima as determined above are shown in Figure 2 as a function of the particle size. It can be seen that the magnetic field needed to induce magnetic

1-

0,7-

0,4-

0,1-

FIGURE 2. The effect of particle radius a on a ratio Sf/B£ of threshold magnetic fields for floccutation in primary and secondary minima. 1 A tf0 = 30mV, K = If/rn" 1 B $0 = 30mY, K = lO'm" The curves are independent of the magnetic susceptibility of the interacting particles

39 flocculation into the secondary minimum can be up to an order of magnitude smaller than the field required to suppress a potential barrier. This model of magnetic flocculation of fine, weakly magnetic mineral particles allows one to analyse the flocculation of particles into primary and secondary minima. It follows from the model that, if the particles are not too small and the magnetic induction not too high, a secondary minimum does exist, and, owing to the flocculation into a secondary minimum, a threshold magnetic field required to induce rapid flocculation can be reduced by an order of magnitude. An appreciable reduction of the energy required to generate a magnetic field could be achieved, and it becomes possible even for diamagnetic particles to be flocculated at magnetic fields that are industrially feasible.

Paper A5/3 SINGLE-WIRE MAGNETIC SEPARATION: A DIAGNOSTIC TOOL FOR MINERAL PROCESSING by

J.H.P. Watson and D. Rassi University of Southampton England

Introduction i High-gradient magnetic separation (HGMS) is now widely recognized as a powerful technique in the mineral-processing industry. Pilot studies to determine the suitability and profitability of HGMS for a given mineral sample, however, are expensive and time-consuming. A comparatively simple method for the study of the magnetic properties and behaviour of mineral samples was therefore devised. This involves the capture of the particles under study onto a ferromagnetic wire in an external magnetic field. The material thus captured can be collected and analysed visually (e.g. by use of an electron microscope) or spectroscopically (e.g. by X-ray-fluorescence spectrometry). Furthermore, the shape of the build-up on the wire yields important information on the constituents of the sample. Figure 1 shows the geometry of the single-wire setup. In certain situations it may be advantageous to have the flow perpendicular to the field direction (i.e. flow along the y-axis in Figure 1). In either case there are two zones of paramagmetic capture and two zones of diamagnetic capture on the wire surface. A comprehensive theory of capture onto a wire has been developed by Watson.

Experimental The single-wire cell consists of a Perspex channel incorporating the wire mounted on a plug that screws through the back wall of the channel. The cell is placed between the poles of a magnet with the wire in the region of maximum field, while the sample to be studied is passed through the channel at a controlled rate. An optical window

40 FIGURE 1. Basic filter configuration. A ferromagnetic wire of radius, a, placed axially along the zaxls in a uniform magnetic field, Ho, applied in the xdirection interacts with a paramagnetic particle of radium, R, in a moving field. The fluid flows past the wire with a velocity, Vo, in the negative, x, direction is placed in front of the wire, through which the wire cross-section can be viewed by means of a microscope. The image can then be projected on a television screen by a closed-circuit television system. Alternatively, a polaroid camera can be mounted on the microscope to take still photographs of the build-up of particles on the wire. A schematic diagram of the single-wire apparatus is given in Figure 2. • Video Video recorder monitor

Electromagnet

^.Cell ^^ LJlAy Video V — camera P P"NJJ U O source If \ A ' •JIA IK /v v Microscope . ... >-v 1 "SGollection ' 1 Electromagnet 1 \ • FIGURE 2. Schematic diagram of the single-wire apparatus

\ 41 . I

When a visual inspection of the particle trajectories and build-up is completed, the captured material is collected for further analysis by the introduction of a hypodermic needle into the channel through a rubber diaphragm on the side wall of the channel. The needle is then operated to suck the deposited material off the wire.

Results and Discussion As mentioned earlier, the shape of the build-up on the wire is indicative of the type of capture. Ferromagnetic particles, for example, form long chains, and the deposit has a spiky appearance. Paramagnetic material, by contrast, is deposited as two smooth lobes in the paramagnetic capture zones on the wire. In a recent investigation of the treatment of gold-ore leached residues by HGMS, the sample was studied in the single-wire cell by the method outlined above. The presence of a small (approximately 4 per cent) ferromagnetic component was thus discovered, and its removal led to much improved beneficiation of uranium and gold.

Paper A5/4 BACTERIAL HEAP LEACHING OF LOW-GRADE NICKEL MATERIAL

by

P.C. Miller, I.J. Corrans, and A.J. Southwood Council for Mineral Technology South Africa For many years, heap leaching has been recognized as a viable economic process for the extraction of metals from ores. It has been used successfully for the extraction of copper, gold, and uranium from ores that are too low in grade to be considered for treatment by conventional metallurgical routes. Although it is one of the oldest and conceptually simplest methods of metal extraction, it has received little attention in Southern Africa. An extensive research programme has started at the Council for Mineral Technology (Mintek) on the possible application of this method to large quantities of low-grade material in Southern Africa. Typically, this includes the treatment of existing dump material; or the exploitation of large virgin ore-bodies by heap leaching of the crushed ore or of a low-grade concentrate produced by various ore-dressing techniques. Particular emphasis is currently being placed upon the recovery of nickel from low- grade material, and the paper reports on the progress of the present study, drawing extensively from the results obtained in laboratory experiments. Although nickel is the prime metal of interest (present as pentlandite-pyrrhotite), both the concentrates and the ores used in the experiments contain significant quantities of cobalt and copper (chalcopyrite). The advantageous presence of the bacterium Thiobacillus ferrooxidans has been demonstrated both in shake-flask tests and column tests. In 20kg column tests, the recovery of nickel from sulphide material was over 20 per cent higher in the bacterial leach than in the sterile control. The bacterial leach released 10 per cent of the total

42 nickel in 160 days from ore sized between 6 and 12mm. Both nickel and cobalt were i leached preferentially to copper. | 'Mini'-column tests were established to show the leaching characteristics of ores and j dump material that are potential materials for heap leaching. The columns contain '• approximately 1,5 kg of material sized between 0,5 and 1,7 mm, and are irrigated with i a weak acid solution containing bacteria. Generally, the recoveries of nickel and cobalt : vary from one ore to die other, as they do from different dump materials. Such variations are closely associated with the mineralogical nature of the material concerned. The recoveries of nickel varied between 1,7 and 6,3 per cent for the ores during a leaching period of 130 days, and a recovery of 38 per cent was obtained from a dump material. Three 20 kg column tests using ore of different particle sizes were conducted to illustrate the importance of particle size on leaching. The results indicate that, in the initial leaching stage, the effect of particle size on leaching can be misleading in the presence of a bacterial lixiviant containing low concentrations of ferric ions. The initial recoveries of nickel sulphide in which the exposed mineral is attacked, were controlled by the availability of ferric ions, and not by particle size. As the concentration of ferric ions increases in solution with the leaching time, and the availability of exposed metal 1 minerals decreases, the rate of recovery is likely to be controlled by the rate of diffusion I of ferric species and products through the rock matrix, rather than by chemical-reaction rates. Particle size therefore becomes increasingly important as leaching progresses, unless the rock matrix is readily degradable. The methods under consideration for the production, from an ore, of a low-grade concentrate suitable for heap leaching include combinations of the following: ore sorting, magnetic separation, gravity separation, heavy-medium separation, and flotation. As , many of these methods give a product of fine size, or a large amount of fines are produced, it is necessary for the material to be agglomerated prior to heap leaching. Two 20 kg columns of an agglomerated low-grade nickel concentrate were established. One contained agglomerates obtained merely by the crushing of an already fused pyrrhotite concentrate to between 6 and 12 mm, while the other column contained agglomerates produced by the binding together with sand and cement of the fine material • from the same concentrate. Within 60 days, 45 per cent of the nickel had been leached from the 'natural' agglomerate, whereas only 18 per cent had been recovered from the 'man-made' agglomerate. The importance of agglomeration is discussed in the paper. When a column containing 92 kg of unagglomeraled low-grade concentrate was leached, a recovery of 30 per cent was achieved within 70 days. In the absence of oxygen, the production of hydrogen sulphide (a degradation product of pyrrhotite) gave serious problems because nickel precipitated from the solution. Practical methods of alleviating such a problem include aeration of heaps and the use of thin-layer leaching. An example of how heap leaching of a concentrate can be more economical than ! '< heap leaching of crushed virgin ore is given in the paper.

43 4 SESSION A6 . Innovations I Paper A6/1 r j THE DESIGN AND CONSTRUCTION OF A SUPERCONDUCTING MAGNET | FOR PRODUCTION-SCALE DRY SEPARATION OF MINERALS

by

J.A. Good and K. White Cryogenic Consultants Ltd England The development of superconductivity has made available significantly higher magnetic fields over larger volumes than previously available. Further, the cost of producing even modest magnetic fields over large volumes has been greatly reduced. This is made possible because superconducting windings carry large electric currents at high current ! densities without any power dissipation. A superconducting magnet requires no iron, ' _ and is therefore not limited by the saturation magnetic flux of iron, which limits conventional resistive electromagnets to maximum fields of about 2 T. The current density that can be achieved in a superconducting winding is a function of the magnetic field. For niobium-titanium superconductors, a practical figure for current density in a winding is 4 x 10* A/cm2 in a magnetic field of 4T. If attempts are made to increase the current in the winding above the superconductor's maximum \ current-carrying capacity, the superconductor undergoes a transition to a highly resistive state, and the current and magnetic field are extinguished. This phenomenon is known as quenching of the magnet. For niobium-titanium alloys, the maximum permitted current density of 4 x 10* A/cm2 reduces linearly to zero as a function of field from 4 to 10 T. Although it is practical to make superconducting solenoids with fields of 8T from niobium-titanium alloy, the cost is greatly increased because the volume of windings, and therefore of conductor, is greatly increased. Superconducting windings are best suited to the generation of relatively homogeneous magnetic fields. For magnetic separation, a strongly divergent magnetic field is required. There are two methods by which this can be obtained. One method involves the use of shaped-iron pole pieces as magnetic flux concentrators, which produce very high, short-range magnetic forces. Separators of this type are widely used in the minerals industry with conventional resistive electromagnets. The Jones and Sala magnetic separators are examples. ,, In order to obtain strong magnetic-field gradients using superconducting windings, • it is necessary to design electromagnets with strongly divergent field shapes. For the »\ MK III prototype magnetic separators developed by Cryogenic Consultants Ltd (CCL), a reverse pair of windings is used. Two circular coils, each of 355 mm external diameter, are placed one above the other in a cylindrical cryostat. The cryostat is a stainless-steel vessel containing all the thermal insulation and refrigeration equipment necessary to maintain the superconducting coils at their operating temperature of 5 K. The ov'.tr diameter of the vessel is 365 mm, and the magnetic field at the surface of this cylinder

44 is 3,2 T. The field falls away with distance from the cylinder, with a gradient of 0,8 T/cm. The separation system developed around the magnet is described elsewhere in the two accompanying papers. Ore is passed down the surface of the cryostat while it is standing with its axis vertical. The ore is separated into two streams with a knife-edge splitter. For the MK III machine, all the equipment for controlling the flow of ore and separating it has to be developed to fit onto the curved outer surface of the cylinder. For a device such as the splitter, this is a serious complication. The specification called for by Foskor for its new pilot plant required a separator to handle about 601 of feed per hour. The design proposed by CCL and now constructed is based on a linear geometry rather than the earlier circular windings. The cryostat and magnet were specifically developed for the dry process that had been developed at Foskor. A powerful dipole superconducting magnet is enclosed in a rectangular-section cr/ostat. The cryostat is 3,5 m long, 450 mm high, and only 85 mm wide. A strongly divergent magnetic field appears on each side of the cryostat. The design calls for an increase in magnetic field strength at comparable gradients, with a very much larger volume than the MK III machine. Separation takes place on both of the two sides, each of which has a magnetic zone 3 m long by some 150 mm high. The mechanism for controlling the ore feed and for separating the mineral and collecting the products can now be constructed to work against a flat surface. Control of the position of the splitter is very much simplified as compared with a circular geometry. While the design of the separation system is much simplified by the linear geometry, the design of the cryogenics and superconducting system is greatly complicated. The refrigerator, which is similar to that used in the MK HI machine, is contained in a cylindrical vessel at one end of the magnet cryostat. The 3 m-long superconducting dipole represented a serious challenge to magnet technology. Very large forces have to be contained to restrain the superconducting windings. An analysis of the principal design problems is presented, together with the results of experimental testwork. Paper A6/2 INDUSTRIAL-SCALE DRY BENEFICIATION OF PHOSPHATE-BEARING PYROXENITE ORE

by

E.H. Roux, J.G. Goodey, E.F. Wepener, and K.R. Hodierne Phosphate Development Corp. Ltd South Africa

I I For the past thirty years the Phosphate Development Corporation has been recovering phosphate minerals by means of flotation from the vast Palabora Igneous Complex. Mineable reserves exceed 21 milliard tons of ore and can supply South Africa's domestic phosphate requirements for at least 1500 years. Current production is twice the domestic demand, and the surplus is exported. Of the ore-body, 95 per cent consists of pyroxenitic ore, in which the phosphate mineral (apatite) is diamagnetic, whilst all the gangue minerals (dolerite, phlogopite,

45 1 hedenbergite, diopside) are paramagnetic. In 1978 laboratory tests on a Frantz i Isodynamic Separator indicated that a high-grade apatite concentrate could be produced | with acceptable recovery by means of dry high-intensity magnetic separation (DHIMS). j A successful DHIMS process on an industrial scale can potentially offer a simpler | and more cost-effective process than flotation by eliminating the use of water, pumps, • pipelines, slimes dams, thickeners, filters, dryers, flotation reagents, flocculants, and fuel for drying. Extensive laboratory tests on induced magnetic-roll (IMR) separators showed that 85 per cent of the pyroxenite ore, after being dry-ground to mineral-liberation size, could be processed efficiently by means of DHIMS. The remainder, being fines smaller than 50 pm, responds better to flotation. Preliminary cost calculations indicated technical feasibility and potentially attractive economics, and a 100 t/h pilot plant was built. This plant was commissioned early in 1982. Since IMR separators are the only well-developed, commercially available dry high- intensity magnetic separators, the pilot plant has been equipped with two lines of these J machines capable of processing 9t of ore per hour in a five-stage separation. Owing I _ to their low individual capacity, the large number of moving parts, the number of feed points, air gaps, and splitters that have to be controlled, their energy consumption, and their heat output, IMR separators are not considered to be suitable for a large ore-beneficiation plant. Alternative separators under investigation include ELB-Yaniv perm-roll separators, the SALA high-gradient matrix type of separator as modified for dry feed, and a jj superconducting magnetic separator of novel design that has been developed in collaboration with Cryogenic Consultants Limited, of London. The relative merits of the various candidate separators are discussed. Results obtained in the pilot plant are presented, and problems encountered in the development of the dry-processing technology are discussed, as well as proposed solutions. The projected economics and operational features of dry processing relative to those of flotation in the case of pyroxenite ore are discussed.

Paper A6/3

FAST FLOTATION WITH AN AIR-SPARGED HYDROCYCLONE

by i J.D. Miller and DJ. Kinneberg : University of Utah U.S.A.

In conventional flotation practice, hydrophobic particles attach to air bubbles and are collected in a froth phase at the top of a stirred tank. Two important factors in determining the effectiveness of the separation are collision frequency and particle size.

46 The collision frequency between hydrophobic particles and suspended air bubbles influences the rate of flotation. With conventional stirred-tank cells, residence times of at least several minutes are required for adequate separations, primarily because of the randomness of particle-bubble encounters. The particle-size limit (i.e. the size below which flotation response decreases significantly) affects the overall recovery for a particular separation. In general, this limit is about 10/un for sulphuric minerals, and as much as 100 pm for non-sulphuric mineral commodities. Consideration of these two limitations, flotation rate and fine-particle flotation, led to the idea of an air-sparged hydrocyclone in the late 1970s. The device was envisioned to take advantage of a controlled high force field developed by cyclone fluid flow, together with a high density of small bubbles, in order to provide directed particle-bubble encounters and thus effective flotation of fine particles at accelerated rates. The basic features of an air-sparged hydrocyclone are (1) a porous wall through which air is sparged; and (2) a tangential flow of shiny orthogonal to the air flow. A number of different designs incorporating these features have been tested, the preferred design being a vertically oriented cylindrical cyclone with a tangential feed at the top (Figure 1).

The stabilizing properties of the froth and the low circula- tion velocities in the centre Fine particles, requiring high of the cyclone create a very i force fields in order to float, quiescent froth column for are injected at the point of hydrophobic particles. maximal centrifugal force

Hydrophilic particles re- Injection of air through the jected with the major por- porous cyclone wall, where tion of the water the shear rate is maximal, ensures the formation of small air bubbles

FIGURE 1. Diagram of an air-sparged hydrocyclone

47 The slurry, fed through a conventional cyclone header, passes through the device as a thin layer and travels downwards countercurrent to the froth phase, which is moving upwards in the centre of the device to an axial discharge. Hydrophilic particles are thrown to the porous wall and are discharged as reject. Hydrophobic particles, also thrown to the wall, encounter radially sparged air bubbles. The high shear force at the wall generates small bubbles and provides intense particle-bubble interaction. Attachment occurs, and the hydrophobic particles are transported to the upward-moving froth phase. Previous publications have discussed the application of air-sparged hydrocydones to the flotation of coal and of copper-porphyry ores. Coal would appear to be ideally suited to this device because its density is lower than that of most gangue constituents. The driving force for flotation is therefore superimposed on die driving force for gravity separation. However, minerals with densities greater than gangue constituents can also be concentrated with an air-sparged hydrocyclone. To demonstrate this fact, a typical western copper-porphyry ore was selected for investigation. Besides the density difference, copper flotation differs from coal flotation in the degree of concentration required for effective separation (head analyses for the valuable constituents of copper- porphyry ores are usually less than 5 per cent by mass as against 30 to 70 per cent for the valuable constituent of coal) and in the strength of bubble-particle interactions.

100 1 Modified ' 1 i air-sparged \§^j hydrocyclone ^S$ I 80 -

g, 60 .- % A - hydrocyclonAir-spargede v\Ayyf < 13 to 16 minutes usual range of 40 -- [Conventional /of industrial flotation ^^£/ roughing retention based on /y times results of yNy s 20 - batch tests] jRy

1 I 1 1 0,1 1,0 10 100 1000 Nominal retention time, seconds FIGURE 2. Comparison of air-sparged hydrocyclone recoveries with those predicted for conventional continuous flotation based on batch tests

48 Thus, copper flotation provides an interesting contrast to coal flotation with regard to the development of the air-sparged hydrocyclone. Froth flotation in the air-sparged hydrocyclone is reviewed with regard to flow behaviour, performance, cost, and design variations. For example, a comparison of die performance of the air-sparged hydrocyclone with conventional flotation for a selected low-grade copper-porphyry ore showed that remarkably short retention times were required to achieve satisfactory separations (Figure 2). Moreover, preliminary data indicated that the recovery of molybdenum and gold with the air-sparged hydrocyclone was comparable with that obtained in a conventional cell. It Is concluded that the air- sparged hydrocyclone has the potential to significantly alter conventional flotation technology, although additional research and development for scale-up will be required before this potential can be realized.

Paper A6/4

SOME ENGINEERING GUIDELINES ON FLOTATION REAGENT USE

by

R.R. Kiimpel Dow Chemical Company and The Pennsylvania State University U.S.A. The industrial practice of separating finely ground valuable minerals from associated gangue materials by froth flotation is well known. Despite the obvious importance of this industrial process, the predictive tools for the plant-engineering evaluation of changes in flotation-process variables are still in a developmental stage. Consider, for example, the situation where an operating engineer is faced with evaluating a change in an existing chemical-reagent scheme, possibly based on batch laboratory data. When all is said and done, the engineer must still perform detailed physical testing of the proposed reagent change on the actual equipment in plant use to be confident, even of the direction of change of plant results, let alone the economic viability of the change. The theory behind the behaviour of chemical reagents in froth flotation has been studied extensively, and, certainly, important progress has been made. Unfortunately, the nature of the flotation process, with its complicated heterogeneous surface-chemistry activity operating on highly variable and difficult-to-characterize ore surfaces, coupled with complex phenomena of mass and momentum transfer, seriously compromises the plant-scale predictive capability of theoretically based models. Faced with this situation, a programme of study was organized by the author in the middle 1970s with the goal of measuring and characterizing, in an organized and consistent manner, a large number of flotation tests performed at different plants:. In particular, it was desired to identify, if possible, some general engineering-oriented guidelines on chemical use. This study involved both batch laboratory and continuous plant-scale data, special emphasis being placed on defining the role of the type and dosage of chemical reagents.

49 In order to achieve useful reagent correlations, it was found necessary to treat flotation as an interactive 'engineering system' as demonstrated in Figure 1. Experience has shown that successul industrial-scale flotation practice involves knowing the capabilities (strengths) of the various system components to compensate for physical or economic limitations (weaknesses) inherent in the particular settings of other components. Thus, skilled operators will, over a period of time, arrive at an 'optimized* flotation operation by balancing the various system components under their control.

Collectors Chemistry Frothers components Activators Depressants PH

Equipment components Operational components Cell design Feed rate Agitation Mineralogy Air flow Particle size Cell-bank configuration Pulp density Cell-bank control Temperature FIGURE 1. The process of flotation illustrated as a three-cornered interactive system One serious problem in current flotation practice is that most often the relative contribution of each of the various variables towards improving a given operation is not known qualitatively, let alone quantitatively. For example, assume a current flotation operation is experiencing a mass-removal rate limitation for a given cell configuration, feed rate, etc. What are the system component settings (changes) that can help improve this situation and, even further, which variable settings from each of the three corners of Figure 1 are more likely to dominate other variable settings in their ability to increase

50 /'••

mass-removal rate? Also, another related question is: what will be the consequences of making one or more changes in the system components on other facets of the flotation operation that currently may be working in a suitable manner? This paper summarizes some of the engineering trends observed in the mineral flotation portion of the study described above.

51 SESSION A7 Modelling Design and Control

Paper A7/1

THE PROCESS ENGINEERING OF SIZE REDUCTION USING BALL MILLS

by

R.R. Klimpel* and L.G. Austinf *Dow Chemical Company tThe Pennsylvania State University U.S.A. The procedure for the scaling of breakage parameters determined in a laboratory mill to values for a full-scale mill is briefly presented. A simulation model of a closed-circuit mill also requires a model of residence-time distribution, a mass-transfer law to predict over-filling as a function of flowrate through the mill, and a model to describe classification action. For fine dry grinding or viscous-slurry grinding, it is necessary to allow for slowing-down of breakage rates for a fine mill charge. Two examples of the comparison of simulation models to actual plant data are presented. The first is wet grinding of copper ore in normal closed circuit at 50 t/h production, under conditions where the mill does not overfill. The second is the wet grinding of copper ore in reversed closed circuit at a production of 87, 102, and 108 t/h, with flow through the mill of I M approximately 150, 200, and 300 t/h. In all cases, the simulated capacities and product size distribution were in reasonable agreement with actual plant performance. The objective of a simulation model for a tumbling mill is to enable the prediction of mill capacity and product size distribution as a function of mill design and operating conditions, with sufficient accuracy for economic costing and precise design specifications. The empirical design method of Bond has been very valuable, and elements of his method are incorporated in the procedure described here. However, it has a number of limitations. To improve upon his method, it is necessary to use the more sophisticated technique involving the solution of the steady-state, continuous grinding equations in closed-circuit configurations. Formulation and solution of the equations in terms of specific rates of breakage (Sj) and primary breakage-distribution functions (Z?jj) are in themselves not sufficient. It is necessary that a description be obtained of the way these parameters vary with mill geometry and mill conditions. This has occupied attention for a number of years and has involved measurements in laboratory-scale mills as a function of ball load (/), powder hold-up expressed as a fraction of interstitial filling (U), rotational speed as a fraction of critical speed (0C), mill diameter (D) and length (L), ball diameter (d), and the mixture of balls in a mill, slurry density, etc. In addition, accurate mill description required the development of a sufficient knowledge of the residence-time distribution of small- and large-scale mills. It has been necessary to analyse and describe classifier performance in terms of partition (Tromp) curves. Finally, a mass-transfer law is

52 39

essential: this describes the accumulation of hold-up (W tons or Y) as flowrate is increased through the mill. The conclusions from studies of these factors have been summarized in a recent book entitled The Process Engineering of Size Reduction (AIME, 1984), by Austin, Klimpel, and Luckie, and will not be repeated here. Instead, two examples are given of the use of the methods in the analysis of several industrial circuits to demonstrate that the methods work with some degree of accuracy, and a discussion of uncertainties and future work is also given.

Paper A7/2

PARTICLE TRAJECTORIES AND CHARGE SHAPES IN CENTRIFUGAL MILLING

by

D.I. Hoyer Council for Mineral Technology South Africa

Paper A7/3

MULTIVARIABLE CONTROL AT FRANK CONCENTRATOR

by ;| I.J. Barker*, S.C. Axcellt, and D.G. Hulbert* •Council for Mineral Technology tRustenburg Platinum Mines South Africa For many years, the Frank Concentrator of Rustenburg Platinum Mines has practised milling control by on-line computer. Recently, in conjunction with the Council for Mineral Technology (Mintek), multivariable control was successfully implemented on the A section of the concentrator. The milling circuit being controlled consists of two primary ball mills and two secondary ball mills closed by hydrocyclones. In addition, there are two unit flotation cells in the secondary circuit. The object of the control is manipulation of the feed to the circuit and of the dilution water to the cyclone feed sump so that the particle size of the product (as measured on-line) and the circulation load are kept at fixed operating points. First of all, the necessary instrumentation had to be installed and debugged. For this purpose an on-line instrument-based mass balance in conjunction with an off-line sample-

53 based mass balance proved invaluable. Next, step tests were conducted on the circuit to produce the dynamic responses, which were logged by the computer and were then analysed by use of the software available at Mintek for this purpose. In this way, a multivariable controller was eventually designed by use of a technique involving Inverse Nyquist Array. This controller was installed on the computer at the plant and seems to be operating satisfactorily. Multivariable control is simply a way in which a plant can be controlled to achieve a desired operating state. Its main direct benefit is that it reduces fluctuations in the particle size of the product, thereby providing a better feed to the flotation plant downstream. However, this is not the only benefit. Mere preparation of a plant for multivariable control produces certain benefits. For example, it was found that the installation of cascade control on the cyclone feed sump resulted in a much more consistent product size. Secondly, by the achievement of a desired operating point on the plant once a controller has been installed, an optimization campaign can be conducted in which bottlenecks can be removed and the operation improved. Such an optimization campaign is currently under way, and some of the initial findings are reported.

Paper A7/4

THE USE OF SIMULATION IN THE OPERATION AND DESIGN OF MINERAL PROCESSING PLANTS

by

DJ. McKee and A.J. Lynch Julius Kruttschnitt Mineral Research Centre Australia

Satisfactory mathematical models exist for many of the common unit processes in mineral I processing, particularly crushing, screening, grinding, classification, and flotation. The J use of these models for simulation studies is well developed. However, most of the actual simulation work is performed by specialists rather than plant metallurgists or plant operators. This is because the models are generally poorly documented, operating instructions are inadequate, and very detailed knowledge of the mathematical structure of the models is necessary to use their* effectively. At the Julius Kruttschnitt Mineral Research Centre in Australia, a major effort is being made to widen the application of simulation techniques to a range of mineral- processing problems. Like many other research centres in this field, the JKMRC has a comprehensive range of process models and also has many years of experience in conducting simulation studies on real industry problems. The objective of current work is to develop a flexible simulator package which can be used by plant metallurgists to solve some of their day-to-day problems. The simulator also has application in the area

54 ] of operator training and process design. The intention in this paper is to describe the J background to the simulator system, the package itself, and finally its expected uses. j A typical problem suitable for simulation analysis involves the selection of cyclone \ classifiers for a grinding circuit. The first task is normally to mass balance the data | obtained from circuit surveys. This step is not always necessary but as mass-balanced : circuit information is of interest to operators, it is invariably performed. There are many mass-balancing packages available which will satisfactorily perform this task, and in the process provide a set of consistent flowrates, sizings, and assays. The second step is to fit or calculate model parameters using the balanced data. In this case, a set of ball-mill and cyclone model parameters are required. The fitting stage is performed by a second computer programme. Invariably, the data formats for the mass-balance and model-fitting programmes are different. This means that the results from the first programme must be manually transferred to the second. After model fitting has been completed, the parameters must be transferred to the actual simulation programme. This may be an automatic procedure, or it may involve manual transfer of data and setting up new data files. The stage has now been reached i where useful simulation studies can commence. Running the simulations to obtain the ! desired answer is an interactive procedure, involving an experienced simulation specialist. Various flowsheet options and operating conditions are tested, and eventually the best answer is obtained. The objective of the JKMRC simulation is to streamline this procedure and to minimize as far as possible the need for computing and modelling expertise on the part of the user. Provision of the capability for inter-programme communication and data , transfer independent of the user is not a difficult task. A second requirement is the provision of a flexible method of data entry and of circuit-flowsheet specification and results presentation. This must be achieved with a set of highly interactive yet simple instructions. After producing a software specification to suit these requirements, a suitable hardware configuration was sought. The JKMRC has particular requirements in this area. Because research projects are conducted at many mine sites, the support hardware had to be physically small and portable, yet capable of running large Fortran programmes. The operator interface is critical for this system, and the state-of-the-art choice was a high- resolution colour graphics terminal. Finally, total hardware costs had to be relatively low, and a ceiling of $(A)20 000 was stipulated. The hardware purchased consisted of the following: DEC MICRO/PDP-11 computer with 128 K core and 10 megabyte Winchester disk and RSX-11 operating system, Hitachi Success personal computer and colour graphics terminal, and ; DEC LA50 printer. ; All operator communication is via the graphics terminal. Circuits are built up from 1 : a set of standard equipment units which are positioned on the screen using a light pen. Flow connections are then specified with the light pen. A separate programme interprets the screen display and sets up an internal flowsheet description which is used by the models. All graphics displays are processed by the Hitachi, while all model calculations and data storage are handled by the MICRO 11.

55 Three versions of the simulator are being developed. The^first contains all the general- purpose data-handling and process models required for the analysis of steady-state mineral-processing operations. The plant metallurgist is the principal user of this system. A similar version is under development for operator training purposes. A set of very simple commands is available to lead an operator through a series of circuit experiments. In this case, the alarm and diagnostic features of the output are stressed. A third type of package will provide a general-purpose dynamic simulation package suitable for flowsheet analysis, particularly at a design stage.

56 SESSION A8 Modelling Design and Control (continued)

Paper A8/1

REGENT ADVANCES IN ON-STREAM ANALYSIS EQUIPMENT

by

A. Toop, L.R. Wilkinson, and G J. Wenk The Australian Mineral Development Laboratories Australia

The paper discusses some recent improvements in the reliability, stability, sensitivity, accuracy, safety and ease of operation, and calibration of the instruments which are used for on-stream analysis in mineral-processing plants. These improvements have mainly been achieved by the incorporation of solid-state devices and microprocessors into the instruments. The benefits of these advances in technology are discussed.

Advances in Technology Some examples of the advances in technology include the introduction of (1) solid-state X-ray detectors for the on-stream analysis of mineral slurries or solutions, (2) xanthate monitor for the on-stream analysis of residual xanthate in a mineral- processing stream, and (3) new-generation density gauges for on-stream applications. The new instruments are rugged and reliable, and are based on a modular design so that faulty components can be isolated and replaced quickly to reduce downtime. Simple standardization techniques are used for fault-finding and periodic testing of the equipment. Hence, the recent advances in the equipment for in-stream analysis have greatly improved their usefulness and ease of operation.

Solid-stale Detector Solid-state detectors have been introduced to on-stream analysis during the past three years, and are working reliably in a number of installations. These are energy-dispersive X-ray detectors with good spectral resolution that can resolve X-rays from heavy metals of consecutive atomic numbers. The typical X-ray resolution of a detector in operation in a mineral-processing plant is 230 to 280 eV full-width half-maximum (FWHM) for copper K X-rays or about 3 to 3,5 per cent relative (FWHM/actual peak energy). This resolution is much better than that obtained from proportional counters or scintillation detectors, and it allows simultaneous multi-element analysis with good sensitivity and selectivity. Experience in the use of these probes shows that they are very stable over long periods, but that they must be specially engineered to maintain this stability in plant operation. The probes for in-stream analysis which incorporate the solid-state detector are most

57 useful for the measurement of low metal contents in tailings streams or for simultaneous multi-element analysis, As the solid-state detectors operate at very low temperatures, a supply of liquid nitrogen is required to maintain the temperature of the detector. This is a disadvantage in their use since a continuous supply of liquid nitrogen is required; however, in many concentrators this presents no problem. To eliminate the need for liquid nitrogen, a probe is being developed for the solid-state detector which uses a thermo-electric cooling device. Xanthate Monitor The xanthate monitor uses molecular-resonance absorption spectroscopy to continuously measure the quantity of residual xanthate in a mineral-processing stream. The device samples clear liquor from the mineral slurry in a flotation cell or conditioning tank, and presents the sample to a flow cell, where its optical absorption at several wavelengths is measured. The concentration of xanthate is calculated directly from these measurements. The xanthate monitor measures residual xanthate levels over the range of 0 to 25 p.p.m., and is used as a process-control instrument for the purpose of minimizing xanthate consumption. The measurement technique corrects for the scaling of the optical flowcell etc., and so can be operated for many weeks before a cell wash or replacement is required.

Digital Density Gauge The digital density gauge can be used in stream or on pipe (for the full stream even in a pipe with an outside diameter of up to 1 m) for the continuous monitoring of pulp density. This new-generation density gauge features the use of a scintillation detector and a microprocessor where the older type of gauge used ionization detectors and analogue electronic circuitry. The use of scintillation detectors means that lower-activity radio-isotope sources are required in any particular application, and the safety from radiation is better because these detectors are much more sensitive for radiation detection. Narrow-beam geometry is used, and this also improves the safety from radiation. The incorporation of a microprocessor allows complete digital processing of the electronic signals from the detector, giving better instrumental stability. Other functions performed by the microprocessor include precise linearization of input signals according to theory, calculation and display of output as either percentage solids or specific-gravity units, automatic compensation for source decay, precise calculation of solid mass flow if an input from a flowmeter is connected, and analogue and digital outputs for interfacing the device to other plant instrumentation. Other features of these gauges include a very simple method of scaling outputs and inputs by entering their numerical values into the memory, and a very simple installation and calibration procedure.

58 Paper A8/2 A RECENT ADVANCE IN ON-STREAM ANALYSIS EQUIPMENT FOR MINERAL CONCENTRATORS AND METAL REFINERIES

by

K. Saarhelo* and D.R. Barker? *Outokumpu Oy, Finland tBateman Process Instrumentation, South Africa

Over the past fifteen years, considerable experience has been gained in the construction and operation of on-stream analyser systems for the base-minerals industry. This paper describes a third-generation on-stream X-ray-fluorescence (XRF) analyser that embodies the superior performance qualities of a large centralized analyser system whilst offering a cost-effective solution for the operator of a small to medium-sized mineral concentrator or hydrometallurgical plant. The sample stream under analysis is irradiated by a low-powdered X-ray tube and analysed by a technique of crystal-based spectrometry. When analysing liquids, sensitivities down to part-per-rnillion levels can be achieved, and down to 10 p.p.m. levels when analysing a slurry. The compactly packaged microprocessor-controlled analyser can rapidly and accurately determine the concentrations of five elements simultaneously in five sample streams. Its range includes all the elements between titanium and uranium. In an application where more than five streams require analysis, the system can be expanded to include up to five analyser-probe modules. The various modules can be distributed throughout a plant because they are of rugged construction, which enables them to be directly sited in the process area of a mineral concentrator or metal refinery. The analyser control system (ACS) has been specifically designed to be user-friendly and suitable for interfacing to a process computer-control system. The operator communicates with the analyser through a video display terminal; the assay information is generated on the CRT screen and also as a hardcopy print-out, or is graphically presented in the form of chart recordings. The main task of the ACS programme includes the control of up to five analyser probe units, selectable assay calculation models, statistical data-bank calculations, reporting (log reports, shift and assay reports), and alarm monitoring. A programme for multivariable regression analysis is available for the calculation of calibration coefficients. This programme can be run on-line without disruption of the analyser's operation. Since November 1981, one of the new analysers has been operating in the harsh process environment of the solution-purification section of Outokumpu's cobalt plant at Kokkola in Finland. The analyser has been used to continuously measure highly corrosive sample streams containing various concentrations of cobalt, nickel, zinc, iron, and copper. Continuous on-stream analysis has proved a major factor in improving the efficiency of the plant's hydrogen sulphide precipitation processes. The capital invested in the

59 46

incorporation of the analyser was recovered in well under a year. This was achieved through savings in the consumption of hydrogen sulphide, better yields of the main metals, savings in other chemicals, and the higher production capacity resulting from an increased throughput in the precipitation processes.

Ili

60 47

SESSION Bl Electrolytic Processes

Paper Bl/1

RECENT DEVELOPMENTS IN THE DESIGN OF ELECTROCHEMICAL CELLS FOR THE RECOVERY OF METALS

by

R. Kammel Technische Universitat Berlin West Germany Lately, several novel cell designs have been successfully applied to the electrolytic recovery of metals from dilute solutions like spent leach liquors or electrolytes, rinsing water, effluents, and spills from metal-winning, metal-refining, metal-working, electroplating, and metal-finishing plants. The usual metal concentrations of only a few grams per litre have to be reduced either to predetermined values, after which the solution can be recycled into the process, or to residual values of a few parts per million or less before being drained into a sewer to meet governmental or local environmental regulations. To meet the challenges of these regulations, new technologies have to be introduced within a short period of time to change from the low-waste technology of today to a non-waste technology, and to improve the overall water economy and waste- water situation. In competition with various physical and chemical methods for the treatment of waste water, the electrolytic recovery of metals is steadily increasing in popularity because it is a one-step process, it is clean, and it does not create further waste. In conventional electrolytic cells with vertical electrode plates, the throughput per unit of cell volume is limited not only by the small surface area but also by the poor mass-transfer conditions. Therefore, metal recoveries from dilute solutions can be performed only at low current efficiencies, high energy costs, and insufficient space- time yields. As a solution to these problems, two-dimensional, semi-three-dimensional, and three-dimensional electrolytic systems with electrodes of larger surface have been developed to minimize the current density. To improve the mass transfer and reduce the thickness of the cathodic diffusion layer, the following types of agitation are applied: high-flow turbulent motion, air injection or vigorous agitation of the electrolyte, and movement of the electrodes by vibration, rotation, or mechanical or hydrodynamic agitation. The cell types used by industry at present for the electrolytic treatment of waste water can be classified into concentrator units and direct metal-recovery systems. The concentrator cells operate in such a way as to produce a concentrated solution (electrodialysis) or metal deposits on cathodes from which the metal must be removed periodically by mechanical stripping, by chemical dissolution with oxidizing reagents, or by electrolysis (with fixed-bed cells of the porous, packed, or filament types). In direct-recovery cells equipped with rotating cylinders, fluidized beds, rotating

61 particle beds, and impact-rod cathodes, metal can be recovered in a directly re-usable form, either discontinuously as coherent deposits or continuously as powder, flakes, dendrites, or particles. Recent improvements towards more efficient and intensive process techniques have significantly increased the competitiveness of electrolytic methods of metal recovery, and have extended their application.

Paper Bl/2

THE ELECTRODEPOSITION OF MANGANESE DIOXIDE: THEORY AND PRACTICE

by

R.L. Paul and A. Cartwright Council for Mineral Technology South Africa

During the pyrometallurgical reduction of manganese ores to ferromanganese, a significant proportion of the fines in the feed are blown out of the furnace by the furnace off-gases. The fines undergo partial reduction in the presence of the hot gases before being quenched in the gas-scrubbing system. The resulting 'furnace sludge' has a manganese content of between 20 and 40 per cent, of which approximately 80 per cent is soluble in dilute sulphuric acid. The leach liquor, after being purified, could be used for the production of electrolytic manganese dioxide (EMD). After the feasibility of such a process had been demonstrated, a research programme was initiated at the Council for Mineral Technology (Mintek) on the various unit operations associated with the process. The results obtained for the leaching, filtration, and purification of the manganous sulphate electrolyte are given in a separate paper. A detailed description of a two-part study into the factors affecting the electrodeposition of manganese dioxide onto graphite and titanium anodes is given in this paper. In the first part of the investigation, an electrowinning cell of 5-litre capacity containing three banks of cathodes and two banks of anodes was operated continuously so that the important electowinning parameters could be identified. Purified electrolyte, obtained by the leaching of furnace sludge, was pumped into the cell to maintain the concentrations of manganese and sulphuric acid in the cell electrolyte at the desired steady-state values (approximately 180 g of EMD were produced daily). For each set of experimental conditions, the cell voltage, current efficiency, deposit morphology, and product quality (i.e. impurity content, contents of total manganese and available manganese dioxide, and battery activity) were recorded. The experimental parameters investigated included the concentrations of manganese, sulphuric acid, and impurities in the electrolyte, the temperature of the electrolyte, the current density, the anode material, and the configuration and surface preparation of the anodes. The most important of the electrolyte parameters is the concentration of

62 sulphuric acid, which has a major effect both upon the chemical composition of the EMD and upon the morphology of the deposit. Temperatures below about 88 °C produce anode deposits that are cracked and adhere poorly to the anode substrate. The material used for the anodes, and their shape and surface preparation, have a pronounced effect upon the morphology of the deposit. Graphite anodes in the shape of rods are very tolerant to adverse conditions in the electrolyte (i.e. high concentrations of sulphuric acid and low temperatures), giving smooth adherent deposits under all conditions. However, the removal of the EMD from the rods is difficult without a high percentage of broken rods, and without the introduction of approximately 4 per cent (by mass) of carbon into the product. Sandblasted titanium anodes of various shapes passivated within days when used in electrolytes containing sulphuric acid at levels above about 30g/l. When the anodes were coated with ruthenium dioxide prior to deposition, diey could be used satisfactorily for extended periods in electrolytes containing up to 60 g/1 of sulphuric acid. The second aspect of this investigation involved a fundamental study of the mechanism of the deposition process. Current-time transients and measurements of electrode impedances were used in the development of a model for the deposition process. The proposed mechanism involves at least four consecutive steps, the rate-determining step being the diffusion of manganese ions through a layer of intermediate product to the surface of the manganese dioxide, where an exchange of electrons occurs. The intermediate product, which is produced by the oxidation of divalent manganese, is continuously being oxidized to manganese dioxide. Evidence supporting the covering of the growing deposit of manganese dioxide by such an intermediate product is provided by a study of the redox behaviour of the ferric-ferrous couple by the use of rotating ring-disc electrodes made of manganese dioxide and platinum. Various aspects associated with the operating parameters required for the production of EMD are discussed in terms of the proposed mechanistic model.

Paper Bl/3 THE ELECTRODIALYTIC GENERATION OF SULPHURIC ACID AND AMMONIA FROM DISSOLVED AQUEOUS AMMONIUM SULPHATE

by

P. Steel* and B. Verbaanf * Council for Mineral Technology t University of the Witwatersrand South Africa

Many hydrometallurgical operations require the adjustment of pH by the use of an alkali, and the disposal of the resultant dissolved salts can cause problems. The reaction when ammonium hydroxide is used to neutralize sulphuric acid is as follows:

2 NH4OH + H2SO4 = (NH4)2SO4 + 2 H2O. (1)

63 In this paper it is demonstrated that this reaction can be reversed in an electrodialysis cell by the use of an anion-selective membrane located between two suitable electrodes between which a voltage is applied. The following electrode reactions occur:

Cathode: 2 H2O + 2e -* 2 OH" + H2 (2)

+ Anode: H2O -+ 2 H + V2 O2 + 2 e. (3) The sulphate anions migrate through the membrane to the anode compartment, where sulphuric acid is generated, while the ammonium ions remaining in the cathode compartment associate with the hydroxyl ions to form ammonium hydroxide. The membrane further substantially reduces back-migration of the hydrogen ions to the cathode compartment. The back-migration that does occur reduces the current efficiency of the system to some extent. (The term membrane efficiency would be more accurate than current efficiency in this instance.) Experiments conducted at die Council for Mineral Technology (Mintek) under closely controlled conditions used a cell containing a commercially available anion-selective membrane along with a stainless-steel cathode, a lead anode, and aqueous ammonium sulphate. The effects of the following variables on the current efficiency and on the various voltage-drop components in the cell were studied: temperature, current density, catholyte and anolyte composition, flowrate, and gap width. It was found that, when the temperature of the system was increased, the voltage drops were decreased and the current efficiencies increased. An increase in current density was observed to increase the current efficiency to such an extent that associated increases in voltage drops were more than compensated for, and an overall decrease in the consumption of power resulted. A costing exercise based on economically and technically realistic criteria showed that the total cost of a plant designed to treat 15 mVh of effluent containing 45 kg/m3 of dissolved ammonium sulphate would be about Rl 053 000. Just over 95 per cent of the ammonium sulphate would be removed to produce 13281 of ammonia and 38271 of sulphuric acid per year. The value of the generated reagents would be Rl ,42 million per year, and the total operating cost R692 000 per year. It was assumed that the cells would operate at a current density of 12OOA/m2, and that the 267 m2 of membrane required would be replaced every three years. Experimentally determined values of current efficiency and voltage drop were used. A discounted cash flow showed that the sum of the after-tax cash flows (discounted at 15 per cent per annum) over a project "1, life of fifteen years would be about R4 million. Thus, the net present value would be over R3 million. Expressed in simpler terms, the ammonia would be regenerated at a cost of about R675 per ton, which is about 10 per cent cheaper than the price of anhydrous ammonia. The acid would be recovered at a useful concentration at no cost, and the disposal problem of ammonium sulphate would be largely eliminated. The results of a cost-sensitivity study are presented. Methods for the recovery of the dilute ammonium sulphate from the catholyte (so that it can be recycled to a process) are discussed. Suggestions are made for potential improvements to the process and for further applications of the technology by the use of other soluble salts.

64 I Paper Bl/4 THE ELECTROWINNING OF CHROMIUM FROM TRIVALENT CHLORIDE SOLUTIONS

by

M.D. Birkett Council for Mineral Technology South Africa The electrowinning of chromium from solutions of trivalent chloride salts was investigated so that a procedure could be developed for the manufacture of chromium from ferrochromium. The existing process, known as the Marietta process, which is currently operated by Union Carbide Metals, relies on the sulphate leaching of ferrochromium, followed by slow procedures for the production of a trivalent sulphate solution suitable for the electrowinning of chromium. The chloride leaching of ferrochromium followed by solvent extraction has been shown to be relatively easy, and this provides a potential alternative route to the Marietta process. The electrowinning of chromium from trivalent solutions depends firstly upon the reduction of Cr(III)to Cir(II), followed by the reduction of the Cr(II)to chromium metal. In the Marietta process, the Cr(III) obtained from the separation process is in an electrochemically inactive form, and has to be subjected to an 'aging' process (several days at room temperature) in the presence of ammonium salts before the electrochemically active hexa-aquo species is produced. However, trivalent chloride solutions contain species that appear to be electrochemically active, especially in the presence of high concentrations of chloride, without the need for 'aging' or ammonium salts. Initial electrowinning studies were carried out in a cell fitted with an anion-exchange membrane, which enabled the catholyte reactions to be studied independently of those occurring at the anode. Electrolysis was carried out under a variety of conditions, and the optimum conditions found are compared in Table 1 with those of the Marietta process. TABLE 1

Optimum conditions for theelectrowinning of chromium . Chloride process Marietta process Concentration of Cr (total), M 1,7 to 1,9 0,46 Concentration of Cr(II), M 0,8 to 1,0 0,24 Concentration of chloride, M 6,0 to 6,6 - Concentration of ammonia, M - 4,9 pH value 1,5 to 1,7 2,1 to 2,4 Temperature, °C 30 to 35 52 to 54 Current density, mA/cm2 250 to 300 70

65 It can be seen from Table 1 that electrowinning from solutions containing higher concentrations of chromium can be carried out from chloride solutions, which also results in higher concentrations of Cr(ll). Table 1 also shows that a lower pH value is necessary for chloride solutions. Higher pH values were found to produce hydroxides, and lower pH values resulted in a reduction in current efficiency. This points to the need for good pH control by the addition of acid. The higher pH values employed in the Marietta process, coupled with the presence of ammonium salts, allow for easier pH control. A major advantage of electrowinning from chloride solutions is that operation at high current densities is possible, offering the prospect of much lower capital expenditure. The results obtained under these optimum conditions are shown in Table 2, which shows that the current efficiencies for the electrowinning of chromium from chloride solutions are comparable with those obtained by the Marietta process. However, operation at higher current densities results in higher cell voltages, and thus higher consumption of power. However, it should be remembered that these results were obtained in an ion-exchange membrane cell. It is believed that the voltage could be reduced somewhat by efficient cell design.

TABLE 2

Results obtained under optimum conditions

Chloride process Marietta process Current efficiency Cr(Il), % 70 to 80 Current efficiency Cr, % 48 to 60 45 Cell voltage, V 12 to 13 4,2 Power consumption, kW'h/kg 30 to 40 18,5

An unexpected advantage of electrowinning from chloride solutions was found in an investigation of the effects of metal impurities. As shown in Table 3, much higher levels of metal impurities can be tolerated in chloride solution. The concentrations given there did not appear to affect either the current efficiency or the quality of the deposit.

TABLE 3

Tolerances of metal impurities in the catholyte

Metal Chloride process Marietta process p.p.m. p.p.m. Aluminium 680 22 Magnesium 100 20 Iron 100 53

66 Further studies are being conducted, especially on the quality of the deposit, the use of cells fitted with cloth diaphragms in a continuous-flow system, and electrowinning from the solutions obtained from the chloride leaching of ferrochromium. The results of these and other investigations are reported.

67 SESSION B2 Equipment and Modelling

Paper B2/1

KINETIC MODELS FOR THE ADSORPTION OF GOLD ONTO ACTIVATED CARBON

by

J.S.J. van Deventer University of Stellenbosch South Africa This paper compares and evaluates existing kinetic and equilibrium models for the adsorption of gold onto activated carbon. Batch kinetic experiments have been run for longer than 30 hours, using pure potassium aurocyanide and Au-198 as a tracer for the radiometric determination of gold solution tenor. Different initial gold concentrations, carbon particle sizes, stirring rates, and carbon-to-solution ratios have been used. The results show that granular activated carbon takes up to 4 weeks to reach equilibrium with gold. Powdered-carbon isotherms take about 4 days to reach equilibrium. Up to 40 per cent of the adsorptive capacity is reached in the first few hours, but the remaining capacity is utilized very slowly. This type of behaviour can be described by a dual-rate mechanism macropore- micropore adsorption model. It is assumed that transfer of gold from the liquid to the carbon phase is not a rate-determining factor. The external film diffusion can be written as follows:

NAu = kt(c - cs), (1) in which c = solution tenor, cs = solution tenor at the carbon surface, mass flux across film, and kf = external film mass-transfer coefficient. Equating the fluxes at the outer particle surface gives

(2)

in which Ds = surface diffusion coefficient, / = fraction of adsorptive capacity as macropores, pc = apparent carbon-particle density, ^m = carbon phase macropore loading, and r = carbon radial variable. The macropore mass balance is described by a surface diffusion mechanism:

An yDs a / dqm\

68 where

9% (4) dt

in which q, = carbon phase inicropore loading, and k^ inter-regional transfer coefficient. Equations (1) to (4) can then be solved for the case of a general isotherm:

A C« B~TC»n > (5)

in which A, B, and n are empirical constants. The dual-rate model accurately predicts batch rate data. A quadratic driving-forces approximation of equation (3) gives as good a description of adsorption behaviour as the surface diffusion model. By setting/ = 1, the standard single intraparticle kinetic model is obtained. This model is able to describe only the very early data, and the fitted Ds is a function of the time period over which the data are regressed. By the introduction of a pseudo- isotherm, the model prediction is constrained to level off at the same concentration as the experimental data, and a reasonably good fit to the data is obtained. When liquid-phase mass-transfer resistance becomes negligible owing to rigorous agitation, the boundary condition given in equation (1) is eliminated. This dual-rate model still possesses excellent simulation properties. The Langmuir adsorption model uses one rate constant and simulates only the very early rate data. A forced fit can be obtained by using a pseudo-isotherm. Existing empirical rate models, such as the Fleming model, are unable to simulate kinetic behaviour over a wide range of conditions. This work shows that a dual-rate model is required to simulate the adsorption of gold onto activated carbon over extended periods.

Paper B2/2

THE MEASUREMENT OF CARBON CONCENTRATION IN THE ADSORPTION VESSELS OF A CARBON-IN-PULP PLANT

by

N.D. Hulse, B. Fitzgerald, and M.J. Ohlson de Fine Council for Mineral Technology South Africa

It is an established fact that a critical factor in the gold-winning process by carbon-in- pulp (CIP) is the amount of suspended carbon in each reaction vessel of the system:

69 insufficient carbon causes a loss of gold whereas too high a concentration is wasteful, and both extremes are economically unacceptable. These factors provided the motivation for the Council for Mineral Technology (Mintek) to investigate the possibility of developing a system for continuously monitoring the concentration of carbon in each adsorption vessel of a CIP plant. All the experiments leading up to the design of a 'carbon meter' were based on the fact that, when a particle of activated carbon strikes a metallic electrode, an electrical pulse is generated by the fractional discharge of the particle. Details are given of initial laboratory experiments on the effects of variation in stirring speed, pulp density, and electrode size and geometry on the pulse amplitude. Direct-current biasing of the electrode is also discussed. After the laboratory tests had been completed, a prototype instrument and probe-set were installed at a CIP pilot plant to monitor the transfer of carbon from one reaction vessel to another. These results proved to be very encouraging, and the instrument was transferred to the production vessels of a CIP plant. Details of various probe designs, and the method used for the determination of their optimum operating position in each vessel, are discussed. An improved version of the carbon monitor is being evaluated for long-term reliability. This instrument and its calibration and operation are discussed, together with the results obtained so far.

Paper B2/3 AN INVESTIGATION INTO SOME OF THE FACTORS INFLUENCING THE FILTRATION OF METALLURGICAL SLURRIES M by

G.F. Lahoud and W.A.M. te Riele Council for Mineral Technology South Africa

The factors en which the rate of filtration of a slurry depends can be summarized under two major headings. 1. Characteristics of the Slurry (a) The nature of the solid particles, such as their shape, size distribution, rigidity, and type of surface. (b) The properties of the liquid, particularly its viscosity. (c) The response of the slurry to flocculating agents.

2. Characteristics of the Filter Cake (a) The porosity of the formed cake. (b) The pressure drop across the cake. (c) The thickness of the cake. (d) The permeability of the filter medium.

70 In industry, a wide range of equipment is available for the filtration of metallurgical slurries. All the equipment used separates the solid particles from the liquid in which they are suspended by trapping the solids on a medium that is permeable only to the liquid; thus, a growing layer of cake develops, which, when it has attained a certain thickness, is removed from the medium. Filtration is therefore essentially a batch unsteady-state operation. The approach to the filtration of slurries in the work described was to relate the instantaneous rate of cake formation to the properties of the cake, as well as to the concentration of solids in the slurry and the pressure drop across the cake. The experimental results, which were obtained on a laboratory (vacuum) filtration rig using unflocculated solids from a typical uranium plant, show the influence of pressure drop, solids concentration, cake porosity, particle size and distribution, and cake thickness on the average and instantaneous rates of cake formation. It was found that, for slurries with high contents of solids and for a specific constant- pressure differential, the initial rate of cake formation is rapid, resulting in a random distribution of particles in the cake. Therefore, a cake of low porosity is formed and the rate of cake formation decreases sharply. In addition, sufficient time is available for fine particles to migrate into the bed. Therefore, in this type of bed, the average rate of cake formation is slow. For slurries with low concentrations of solids, the initial rate of cake formation is slower, allowing the particles to segregate within the slurry and the coarser particles to report to the bottom layer of the bed. This segregation results in a more porous bed structure and a smaller decrease in the rate of cake formation. Therefore, the average rate of cake formation is high. The average filtration rate for slurry was found to reach a maximum at a particular pressure drop. For pressures lower than the optimum, the rate of cake formation was sufficiently slow to allow finer particles to migrate into the coarse porous bed that was formed initially, and the average rate of cake formation was therefore slower than the maximum. For pressures higher than the optimum, the initial rate of cake formation was rapid; this favours a mixed-bed formation with lower porosity and, hence, a low average rate of cake formation. In addition, predictions made by the use of a mathematical model based on laminar flow through a porous bed are compared with the experimental results for several size fractions. The influence on filtration rate of porosity and specific surface of cake, as well as filter medium, are highlighted.

71 Paper B2/4

NIMCIX ABSORPTION COLUMNS FOR THE RECOVERY OF URANIUM AT VAAL REEFS SOUTH

by i M.A. Ford* and RJ. Kellyf ft 'Council for Mineral Technology j f Vaal Reefs Exploration and Mining Co. Ltd • f South Africa During the 1970s, there was a rapid increase in the demand for uranium. This encouraged the development of a more cost-effective flowsheet for the recovery of uranium from the low-grade deposits on the Witwatersrand. One of the major modifications to the flowsheet was the incorporation of continuous ion exchange (CIX). The CIX contactor developed at the National Institute for Metallurgy (NIM)—as the Council for Mineral Technology (Mintek) was then called—has been installed and successfully operated at several uranium plants. At Vaal Reefs South, a Himsley CIX system was installed initially. However, when the loading columns, which were made of fibre-reinforced plastic failed, it was decided that they should be replaced with stainless- steel NIMCIX contactors. This paper describes the kinetic model used in the design of the loading columns, and the method followed in the determination of the model parameters from the results of laboratory tests. The model predictions are compared with actual plant performance. In addition, the way in which the NIMCIX loading columns were successfully integrated with the Himsley elution system is discussed. Special consideration is given to the introduction of a microprocessor-based control system for the sequencing of events in the loading and elution cycles.

72 SESSION B3 Leaching Processes

Paper B3/1 i , COMPOSITION AND PHASE CHANGES DURING OXIDATIVE ACID r LEACHING REACTIONS

A.R. Burkin Imperial College England It has long been known that in nature one mineral species may be replaced by another as a result of chemical reaction. This kind of behaviour is also found at times during leaching reactions in which a mineral is altered in stages before complete decomposition occurs. Several different kinds of change can then take place. One of the most frequently studied examples is the two-stage reaction of chalcocite under acidic oxidizing conditions. Rapid dissolution of copper occurs in the first stage and, if particulate mineral is used (about 150/un) at low temperatures (about 40 °C), each particle has a uniform composition throughout. A number of different phases occur successively in the range of composition Cu2S to Cu,S, most of them having a range of stoichiometry, including ranges of composition in which no single phase normally exists. During the second stage of the reaction, CuS reacts relatively slowly, copper dissolving and sulphur being produced, and residual covellite retaining the composition CuS. |3-chalcopyrite of composition CuFeSi83 reacts much more rapidly on acid leaching than does the natural a-form, losing copper, iron, and sulphur in the early stages, although at different rates, so that the copper-iron-sulphur ratio in the residual solid changes. Particles of size about 150/am had uniform composition after partial leaching, and it was found that, as the metals were removed from the sulphur anion lattice framework, certain rf-spacings became smaller, and the lattice collapsed when about 17,5 per cent of the copper had been lost. The lattice spacings then increased again somewhat as leaching continued until, when 31,5 per cent of the copper, 19 per cent of the iron, and 15 per cent of the sulphur initially present in the solid had dissolved (the latter as sulphate ions), a solid composition and structure were reached which did not change further during the remainder of the reaction. The case of acid leaching of bornite, Cu5FeS4, is interesting because it is possible to relate the changes which occur in the rate of leaching as the reaction proceeds to events taking place in the crystal lattice. Copper occurs in two different kinds of site within the lattice framework made from the sulphur atoms; 40 per cent of the copper is in planes parallel to (lll)rh and the atoms can move in these layers unhindered by the presence of any other kind of atom. This explains the rapid first stage of the reaction in which idaite, Cu3FeS+, is formed. Bornite containing 1,2 per cent (by mass) silver replacing some copper was synthesized. This substitution had no clearly significant effect on the rate at which copper

73 ,1 was leached from the mineral, but very little silver was detected in the solution. It was found by examining leach residues that silver which had dissolved from the bornite was cemented out again, taking up positions in a not very well crystalline idaite lattice. Synthetic stromeyerite, CU|O7AgO93S, was leached over the same range of conditions that silver-containing bornite had been. Its behaviour was complex, but there was no gradual change in stoichiometry. As leaching proceeded, mckinstryite, Cu0 8Ag, 2S, and jalpaite, Cu045Ag, 55S, formed and then disappeared successively, Ag2S being the final silver-containing phase while covellite, CuS, was also formed. During the first stage of the reaction of stromeyerite, some copper but no sulphur or silver appeared in solution; the sulphur not required by the sequence of reactions of solids containing silver formed covellite; while at was only the copper not required by all the solids which appeared in solution. Nickel and nickel-iron-sulphides behave quite differently from the copper compounds studied. Heazelwoodite, Ni3S2, when leached with acidic ferric chloride formed millerite, (3-NiS, as a new phase at places on the edges of the particles by the time 2 per cent of the nickel had been leached out. Thus, the rate of appearance of nickel in solution was the result of the leaching of both Ni3S2 and of NiS formed from it. Evidence for the presence of Ni7S6 was sought with great care, but was not found. Pentlandite was synthesized and leached in acidic ferric chloride solutions. Both nickel and iron dissolved at the same rate within experimental error, and elemental sulphur and sulphate were formed. No sign of change in lattice dimensions was observed, and no evidence was found for the formation of violarite, Ni2FeS4, under the leaching conditions used. f Paper BS/2

URANIUM-BEARING MINERALS IN WITWATERSRAND ROCKS, AND THEIR BEHAVIOUR DURING LEACHING

by

G. Smits Council for Mineral Technology South Africa

Uranium is mined and extracted from Witwatersrand sediments as a low-cost byproduct of gold. The recovery of uranium is less satisfactory than that of gold from the same ore, and, although the ores are essentially of a similar nature, recovery varies considerably from one gold mine to another, ranging from 60 to 90 per cent. Consequently, the amount of uranium remaining in the tailings can be substantial. The uranium-bearing minerals that occur in Witwatersrand ores are (1) uraninite (UO2) frequently enclosed by kerogen, (2) 'brannerite'-type minerals (U, _xTi2+JtO6) that formed after the sediments had been deposited, (3) cofllnite (^SiO^i.^OH)^), and (4) uraniferous zircon (ZrSiO4). Their behaviour during leaching was studied at ambient temperature and at 40 and

74 60 °C in polished sections of samples of whole rock from the Witwatersrand, of a ! concentrate of uraninite from the Witwatersrand, and of a brannerite sample of primary ] , origin. The reaction of the uranium-bearing minerals during leaching was studied at > regular intervals under the microscope. i ' It was found that, at room temperature, the dissolution of uraninite started soon after I the sample had been contacted with acid solution, but the rate of dissolution was 1 accelerated by a rise in temperature. Under the correct chemical conditions, dissolution ' was complete within three hours, provided the lixiviant had access to the mineral. However, the 'brannerite'-type minerals (the uranium content of which ranged from trace amounts up to 45 per cent) were less amenable to leaching because they did not react visibly at ambient temperature. The 'brannerite' species of low uranium content were the first to react, and started to dissolve at 40 °C, but those approaching the composition of actual brannerite (UTi2O6) required more severe leaching conditions, dissolution proceeding at a much lower rate. During leaching, the minute inclusions of galena (a product of the radioactive decay of uranium), which form an integral part of uraninite and, in lesser amounts, also of brannerite, reacted with the acid solution to precipitate angiesite (a lead sulphate), whereas the residual product of the various | * brannerite species was titanium oxide, which formed a shielding crust on the surface — of the grains, delaying further dissolution of the uranium from the underlying brannerite. The submicroscopic habit of the 'brannerite'-type minerals (the crystal size being about 10/i.m) and their mantling, and the mantling of uraninite, by minerals that are inert with respect to the lixiviant, are additional factors that reduce the access of leaching solution to uranium minerals, and therefore also play an important part in the incomplete recovery of uranium on several Witwatersrand gold mines. The uranium silicate, coffinite, which is strongly pigmented by galena and frequently enclosed by kerogen, is a rare mineral on most of the gold mines, but is relatively common on those of the West Rand Goldfield. Its behaviour during leaching is not well known, but in the tests it reacted slowly with the acid solution. In the uraniferous zircon, only the metamict portions are uraniferous, and the contribution of uranium by this mineral is negligible. Grains occurring in the leached polished sections showed that the metamict parts reacted, but no detailed information is available on the nature of the reaction.

75 Paper B3/3 THE MODE OF OCCURRENCE OF GOLD AND SILVER IN DOMINION REEF AND THEIR RESPONSE TO CYANIDATION AFTER PRESSURE LEACHING

by

G.W. Glatthaar and C.E. Feather Anglo American Research Laboratories South Africa Dominion Group conglomerates are exploited for uranium, gold, and silver, the uranium being extracted first. Because of the refractory nature of the uranium-bearing minerals, a high temperature-high pressure leach is recommended. While this pressure leach is beneficial to the subsequent cyanidation of gold, a large proportion of the silver present is rendered refractory. By using optical and electron-microscopic methods, the ore and a selection of products from the pressure-leaching plant were examined in detail in an attempt to find reasons for the poor recovery of silver. Gold and silver in the Upper Reef of the Dominion Group in the Afrikander Lease area occur in a number of minerals. Native gold, elect rum, and amalgam are the main gold bearers, whereas silver, in addition to being present in the above alloys, is also represented by native silver, mercurian silver (arquerite), acanthite, stromeyerite, a bismuth-silver sulphide (schabachite or pavionite), and in solid solution in galena. Only minute quantities of these silver-bearing minerals were encountered, and attempts to evaluate their relative abundances in the ore proved to be difficult. It was- possible, however, by use of an electron microprobe to quantify the silver contents in electrum, amalgam, mercurian silver, and galena. Mass-balance studies based on the calculated galena content suggest that about 50 per cent of the total silver is associated with galena in solid solution. Poor silver recoveries are believed to be due to the close association of silver and galena. After pressure leaching, the proportions of both lead and silver increase in the slimes fractions (material smaller than 10 /tin). Prior to pressure leaching, the slimes contain 25 per cent of the silver present. After leaching, the proportion increases to 60 per cent, with a sympathetic increase in lead content. Notwithstanding a careful search, the mode of occurrence of silver in the slimes fractions could not be established. It was noted, however, that jarosite (potassium-iron-sulphate) formed during pressure leaching by a breakdown of sulphides and clay minerals. It is believed that silver is released during the dissolution of the galena, acanthite, other silver sulphides, and native silver, and subsequently precipitated. It may be precipitated as an insoluble complex silver-iron-sulphate (argentojarosite), or may be in solid solution in jarosite. Both compounds are not amenable to cyanidation. A mixture of plumbojarosite (lead-iron- sulphate) and jarosite were also seen to form protective coatings on galena particles, and may occlude other silver-bearing minerals in the same manner. In contrast, finely divided gold particles are liberated from pyrite and other minerals during pressure leaching, and become readily available for cyanidation. The study also suggests that, even if the host minerals are not completely dissolved, a substantial proportion of the gold is exposed to cyanidation.

76 Paper B3/4 OQ £ 4

MODELLING OF THE HARTEBEESTFONTEIN GOLD MINE URANIUM LEACHING AND ION EXCHANGE PROCESSES AND ITS ROLE IN ECONOMIC PLANT OPERATION

by

B.R. Broekman and B. Ward Hartebeestfontein Gold Mining Co. Ltd South Africa Uranium leaching at Hartebeestfontein is performed prior to gold leaching. After milling and gold-uranium-pyrite flotation, the flotation tailings are further milled, thickened, and then dewatered using rotary-drum filters. The dewatered filter cake is dien repulped with an acidic ferric sulphate solution and leached on a continuous basis. Total leach nominal residence time is very short, being less than 10 hours. After leaching, the uranium-rich leach liquor is separated from the solid residue, using double-stage drum filtration, and then subjected to sand and leaf clarification. The clarified pregnant solution then undergoes a process of fixed-bed ion exchange, where the uranium is concentrated prior to precipitation as ammonium diuranate. A series of empirical mathematical models have been developed which satisfactorily describe the major uranium unit operations. Each model was developed using non- linear regression techniques. Where possible, the models were developed having a structure similar to that expected from fundamental considerations. The most significant model developed was that for the determination of the electromotive force of a leach liquor. The model, based on the Nernst equation, is capable of predicting to a standard error of 5,8 mV in the range 350 to 500 mV. The independent variables in the model are temperature, sulphuric acid concentration, and ferric and ferrous concentrations. The uranium-leaching model predicts the leach uranium residue to a standard error of 3,1 p.p.m. in the range 35 to 70 p.p.m. Independent variables include uranium head grade, absolute ferric concentration, electromotive force, and pulp temperature. The model for resin uranium loading indicates that uranium loading is critically dependent upon the resin age and the concentration of several species contained in the pregnant solution. These species include sulphuric acid, nitrates, ferric, and uranium. This model can predict resin uranium loading to a standard error of 7,9kg per charge in the range 100 to 190kg per charge. The above models were used to predict a more economic operating strategy, which, when implemented, would result in a net extra profit of 1,79 million rands per annum. Implementation of the new economic operating strategy required the following changes in operating policy. (1) The leach density was to be increased to 1,7 kg/1. This was initially thought impracticable owing to considerations of pulp viscosity, but laboratory and subsequent plant testwork indicated the feasibility of such high densities. Consequences of the increased leach density include an improved uranium dissolution, a reduction in acid, lime, flocculant, steam and ferric consumptions,

77 and a lower flowrate of pregnant solution, resulting in considerably lower uranium losses in the barren tailings. In particular, the total flocculant consumption for the uranium plant was more than halved to 16 g per ton treated. (2) The leaching temperature was to be increased to 72 °C, which would result in improved uranium dissolution, the value of which would far exceed the cost to increase the temperature. (3) The concentrations of leaching reagents were to be reduced to levels giving the maximum saving in reagent costs with a minimal loss of uranium dissolution, yet still providing conditions necessary for good reverse-leaching properties. The economic use of manganese dioxide as a leaching aid was found to be critically dependent on the price of manganese, sulphuric acid, and lime in relation to the uranium price. A fluctuation in any one of these prices could make the use of manganese dioxide profitable or not. A comparison of the metallurgical accounting figures for the period prior to the implementation of the economic operating strategy with those for the period after implementation indicates that a net extra profit of 1,7 million rands per annum has been achieved. This is within 5 per cent of the predicted net extra profit.

78 SESSION B4 Leaching Processes (continued)

Paper B4/1

LABORATORY TESTING OF ULUDAG SCHEELITE CONCENTRATE FOR THE PRODUCTION OF AMMONIUM PARATUNGSTATE

by

Y.A. Topkaya and H. Eric Middle East Technical University Turkey Laboratory testing of Uludag scheelite concentrate was carried out for the determination of the production parameter of ammonium paratungstate (APT). Autoclave leaching with sodium carbonate, which is best suited to low-grade scheelite concentrates, was chosen as the decomposition technique. After the determination of the optimum parameters of pressure leaching, the purification of the solution and the recovery of tungsten from the leach liquor by solvent extraction were studied. The flotation concentrate which was supplied for the testing contained 24,5 per cent tungsten trioxide and was produced by a flotation pilot plant operated by Etibank at Uludag Tungsten Concentrator. The laboratory screen analysis of the concentrate showed that 83 per cent of the particles were less than 75 fan. Mineralogical investigation indicated the presence of scheelite, calcite, pyrite, apatite, and fluorite, together with smaller amounts of pyrrhotite, magnetite, wolframite, sphalerite, garnet, quartz, etc. •il Some organic material originating from flotation chemicals was also observed. Almost all of the minerals were found to be liberated by the relatively fine grinding used in the flotation pilot plant for the production of Uludag scheelite concentrate. In the decomposition stage, the pressure leaching of the flotation concentrate was carried out in an Autoclave Engineers 1-litre autoclave. The mass of sample for leaching was taken to be 200 g, and the amount of water added to the autoclave was adjusted to give about 100 g of tungsten trioxide per litre in the pregnant solution. The experimental variables investigated were effect of roasting, leaching temperature and pressure, amount of sodium carbonate, and leaching time on the recovery of tungsten trioxide. Laboratory testing showed that roasting of flotation concentrate at about 600 °C for 3 to 4 hours in air before leaching eliminated the flotation chemicals, and prevented foaming and contamination of the leach liquors. As a result, experiments were conducted on the pressure leaching of roasted Uludag scheelite concentrate. The amount of sodium carbonate was varied between 3 to 5 times the stoichiometric amount. The temperature of leaching was in the range 185 to 215 °C, and the digestion time varied up to 5 hours. The optimum conditions were found to be 4 times the stoichiometric amount of sodium carbonate (1,83 kg of sodium carbonate per kilogram of tungsten trioxide) at 205 °C for 3 to 4 hours to obtain a tungsten recovery over 96 per cent. If the soda ash concentration was increased to 5 times, or decreased below about 3 times, the amount demanded by the stoichiometry, the recovery fell. However, by using higher temperatures and pressures of leaching, the amount of sodium carbonate could be decreased to lower levels.

79 After the determination of the optimum conditions of pressure leaching for the roasted Uludag flotation concentrate, stock leach liquor was prepared under the same conditions, i.e. the same autoclave was used several times and the pregnant solutions were accumulated. Later, the removal of silica from this stock leach liquor by magnesium chloride and alum, and of molybdenum by sodium sulphide, was studied. By using these reagents, the pregnant solution, containing 97 g/1 of tungsten trioxide, 460p.p.m. of molybdenum, and 26 p.p.m. of silicon could be purified by precipitation of the impurities. Depending on the experimentally used precipitation conditions and the amount of reagent used, silicon could be reduced to 7 p.p.m. and molybdenum to 25 p.p. m. In the final part of this work, solvent extraction of tungsten from purified stock leach liquor was done with an extractant containing 10 per cent Alamine 336, 10 per cent isodecanol or TBP, and 80 per cent kerosene. Solvent-extraction studies showed that almost all the tungsten could be extracted to the organic phase at 50 °C within 5 minutes at a pH range of 2,0 to 2,5. After washing the loaded organic with deionized water at 50 °C for 5 minutes, the stripping of tungsten was done with twice the stoichiometric amount of ammonia solution at 60 °C for 5 minutes. The strip solution containing about 100 g/1 of tungsten trioxide was later neutralized with slow additions of hydrochloric acid, and the APT crystals that formed on standing for 24 hours were filtered and washed before chemical analysis. The results of this laboratory test indicated that APT could be produced from Uludag scheelite concentrate obtained by flotation.

Paper B4/2

THE DIRECT LEACHING OF OXIDIC MANGANESE ORES

by

W.A. Nattrass and W.A.M. te Riele Council for Mineral Technology South Africa The production, by a process developed at the Council for Mineral Technology (Mintek), of electrolytic manganese dioxide (EMD) for use in dry-cell batteries requires a feedstock of manganous sulphate solution to an electrolytic cell operating at manganese and sulphuric acid concentrations of about 65 g/1 and 40 g/1 respectively. This feed solution is obtained from spent electrolyte by the replacement of its acid content with manganese. Conventional processes use naturally occurring manganese carbonate (rhodocrosite) or manganous oxide obtained by the high-temperature reduction roasting of manganese oxidic ores. The Mintek process is aimed at the achievement of an equivalent result by the direct leaching of manganese ore without a reduction roast. The ore used in the experimental work was supplied by Samancor from their Wessels deposit and was largely braunite (MnO * 2Mn2O3 * SiO2) with some hausmannite

80 (MnO*Mn2O3). If these minerals are regarded as 4MnO*3MnO2'SiO2 and 2MnO*MnO2 respectively, then four-sevenths (about 57 per cent) and two-thirds (about 67 per cent) of their total manganese content can be expected to be soluble in sulphuric acid. The remaining manganese would need to be extracted by reduction leaching. A preferred reducing agent for this purpose is ferromanganese fines, which itself contributes to replenishing the manganese content of the spent electrolyte. The metal content reacts indirectly via a ferric-ferrous couple, the source of iron being partly the ore itself and partly the ferromanganese. The efficiency of the reduction depends on the extent to which the ferromanganese reduces the ferric iron in solution to the ferrous state (which in turn reduces the manganese dioxide equivalent in the ore) compared with the non-productive competing reaction in which hydrogen gas is evolved. The direct evolution of hydrogen from the solution can be minimized by an increase in its content of ferric iron prior to the addition of the reducing agent. One way in which this can be achieved is the use of metallic iron as the reducing agent wholly or together with ferroma>.;ganese. Ferric iron can be derived by the redissolution of iron from die leach residue by the use of make-up acid (required because of the calcite content of the ore), the acid lost as mist during the electrolysis, the 'drag out' in the EMD residues, and the sulphate rejected from the circuit as jarosite. The iron in solution must be eliminated before the manganese-enriched solution can be electrolysed. This is accomplished as follows: all the iron is brought to the ferric state, and the concentration of acid is reduced sufficiently for hydrolysis to occur (at a pH value of about 1,5 to 3,5), when mixed hydroxide species (goethite and jarosite) are precipitated. The leaching process can be considered to involve three main steps: non-reduction leaching, reduction leaching, and neutralization of the leach liquor. The efficiency with which these steps can be carried out directly affects the overall economics of the process. The overall economics depend on the following factors: (1) suppression of the non-productive evolution of hydrogen during reduction leaching; (2) the degree of manganese extraction in non-reduction leaching; (3) the utilization of the reduction leach to ensure that the iron content is finally all in the ferric state, which requires either an excess of ore or the feedback of some product EMD; (4) the achievement of sufficiently low final acid concentrations, which requires an excess of ore to an extent determined by the non-reduction leaching at a pH value up to about 3, or the use of a separate neutralizer, e.g. lime; and (5) the relative prices of the materials used, since the process that is the most efficient in extracting manganese may not be the most cost-effective. These factors are not mutually exclusive. In particular, it has been found that the non-reduction leaching of braunite is constrained drastically by the formation of a calcium silicate slime round the reacting particles, but that this passivity is broken down during reduction leaching, enabling further non-reduction leaching to take place. This passivity is not found with hausmannite or, interestingly, with braunite in a chloride leach. Again, while the passivity of braunite can be overcome in simultaneous reduction and non-

81 i . reduction leaching, the passivity hinders the attainment of the low acid values necessary | in the final stages of the leach. ^ In small-scale batches (approximatejy 4 litres), it was found that up to 90 per cent j of the manganese content of Wessels ore can be extracted, with a utilization efficiency | of about 60 per cent of the ferromanganese fines. Final iron concentrations of less than ; 10mg/l were achieved, and EMD conforming to battery quality was produced from the manganese-enriched solutions obtained. I ) Information is presented on the extent and efficiency of the manganese extraction by the available process options singly and in combination. The results are used to illustrate how the optimum economic conditions, based on the relative costs of the available materials, can be predicted.

Paper B4/3 ENHANCED LEACHING KINETICS OF CHALCOPYRITE FROM CuFeS2/C PARTICULATE AGGREGATES BY FERRIC SULPHATE i L by R.Y. Wan*, J.D. Miller*, and G. Simkovichf * University of Utah fThe Pennsylvania State University U.S.A.

During the past decade, research efforts have been directed to the study of the leaching behaviour of chalcopyrite including electrochemical measurements. The importance of passivating reaction product layers and their electronic conductivity has been established. Researchers at the University of Utah suggested that the rate-limiting step for acid-ferric sulphate leaching of chalcopyrite was the transport of electrons through an elemental sulphur layer which adheres tenaciously to the unreacted chalcopyrite particles as leaching proceeds. The very slow parabolic reaction kinetics were predicted from first principles using the Wagner theory of oxidation. This analysis of the reaction rate accounts for the observed high activation energy (20 kcal/mol) and rate independence of ferric sulphate concentration. Several electrochemical studies of chalcopyrite electrodes have been reported. G.W. Warren and M.E. Wadsworth studied the passive and transpassive anodic behaviour of chalcopyrite in acid solutions. Based on current and mass-balance measurements, they found two intermediate phases appeared to form in the sequence CuFeS2 -* S, -» S2. Other researchers, A.J. Parker, R.L. Paul, and G.P. Power, have suggested that the surface film formed on chalcopyrite during oxidation is a metal-deficient v polysulphide, based on electrochemical measurements. H.G. Linge introduced a model in which the initially formed metal-deficient layer remains beneath the sulphur during the oxidation process. T. Biegler and D.A. Swift studied the anodic electrochemistry of chalcopyrite. They found that 86 per cent of the sulphide sulphur was oxidized to the elemental form (some of which, was plastic sulphur) and 14 per cent to sulphate.

82 They suggest that the physical properties of plastic sulphur may allow it to form a sufficiently dense layer to give the observed low rates of leaching. In all electrochemical studies, the results describe the initial reaction kinetics and indicate that initial reaction product films can slow the rate of reaction by slowing the transport of ions from CuFeS2 m by slowing the transport of electrons. Significantly faster CuFeS2 leaching kinetics are achieved from CuFeS2/C paniculate aggregates in the acid-ferric sulphate system. Leaching results for different carbon types are shown in Figure 1 for 41 pm chalcopyrite aggregates (1,3 cm) containing 9 per cent carbon by mass. The effect of carbon particles upon the rate of dissolution of chalcopyrite has been increased by 400 to 600 per cent (after 10 hours' leaching) depending upon the carbon type which varies in electronic conductivity and carbon particle size. Interestingly, when the carbon content is reduced, even to 2 per cent, the extent of leaching increases further to as much as 7,5 times that achieved in the absence of carbon.

Carbon type Size, fim Surface coverage, % Monarch 800 3,7 74,4 Monarch 1100 3,9 70,3 9,7 28,0 0,6- Spheron 6 Graphite 38 11,9 22,9 Activated carbon 5,3 51,2 No carbon addition 0,5-

CuFeS2 : C = 10 : 1 (CuFeS2 41/i 0,4- Fe2(SO4)3 = 0,25 M H2SO4 = 1M s 90 °C, lOOr/min a 0,3- c o

0,2-

0,1-

10 Time, h FIGURE 1. Leaching behavour of CuFeS^C aggregates for different carbon types

The kinetic aspects of leaching CuFeS2/C aggregates are complicated. More than one rate process may be involved. It is important to note that the leaching rate is improved at lower agitation speeds, presumably an indication that aggregate integrity must be preserved. Also, the rate of reaction for the aggregates is highly dependent on the chalcopyrite particle size. The beneficial effect of graphite addition seems to be more pronounced for coarser chalcopyrite particles. Aggregate size seems to be important

83 only for fine chalcopyrite particles (4,5 fim). These aggregates are more compact, and transport in the aggregate under these circumstances offers significant resistance to the reaction rate. For coarse chalcopyrite particles (41 fim), aggregate size does not influence the reaction rate (aggregate pellet size less than 1,3 cm diameter). As in previous research without carbon addition, the rate of reaction was found independent of the ferric ion concentration. However, a significant temperature dependence of the reaction rate was observed. An apparent activation energy of approximately 25kcal/mol (104,73J/mol) was estimated for the ferric sulphate leaching of CuFeS2/C aggregates compared with 20kcal/mol (83,78J/mol) without carbon addition. Photographs taken with a scanning electron microscope of a partially leached CuFeS2/C aggregate show the formation of the elemental sulphur reaction product to be patchy with areas of exposed chalcopyrite surface. These observations differ significantly from the sulphur structure that forms in the absence of carbon, which is a dense, tenacious sulphur layer. The elemental sulphur formed in the presence of carbon seems to wet attached carbon particles, and a large portion of the chalcopyrite surface remains unprotected. These results indicate that the reaction rate is no longer controlled by parabolic kinetics, and other explanations must be offered to explain the experimental results. \ Paper B4/4

THE LEACHING OF BASE METALS FROM THE CALCINES PRODUCED BY THE ROASTING OF PYRITE CONCENTRATES

by

M.J. Nicol* and A.O. Filmert "Council for Mineral Technology, South Africa tFormerly Council for Mineral Technology, South Africa; now CRA Services Ltd, Australia

A number of gold and uranium plants in South Africa concentrate the pyrite in the ore or residue by flotation, and roast the concentrate for the production of sulphuric acid. The calcine produced, which is predominantly hematite, is generally subjected to cyanidation for the recovery of the gold and silver that would normally not have been recovered in high yield from the primary concentrate. The calcines often contain economically significant quantities of copper, nickel, cobalt, and uranium. Prior treatment of the calcine for the recovery of these metals would be desirable in terms of the value of the products, and also in terms of the possible increased gold recovery in the subsequent cyanidation of the leach residue. Several processes for the leaching of the base metals from plant calcines have been investigated, and an important general conclusion is that adequate recovery of the base metals requires that a large proportion of the iron should also be extracted. This observation led to a more extensive investigation of the kinetics of the leaching of various

84 \ iron oxides. The application of electrochemical theory and techniques resulted in a fuller ! understanding of the various factors that govern the rate of leaching of iron oxides. > , As a result of this fundamental work, alternative treatment schemes that should yield ! more efficient extraction from calcines were suggested. Several of these possibilities were ] investigated, and the most promising were found to require reducing conditions during ! the leach, or prior partial reduction of the calcine to magnetite or wustite. The effects of various process variables such as the type and concentration of the acid, the temperature, the residual sulphur in the calcine, and the source of the calcine were investigated. The effect of prior treatment of the calcine on subsequent gold recovery by cyanidation was also investigated for a number of the calcines.

85 k SESSION B5 Separation Processes

Paper B5/1

PRECIPITATION OF METAL VALUES FROM CATIONIC EXTRACTANTS

by

A.J. Monhemius Imperial College England The recovery of metal values by direct precipitation from metal-loaded organic phases in solvent-extraction operations may lead to simplified process flowsheets, since two conventional steps, namely aqueous stripping and metal winning, are combined into one. Furthermore, the substitution of an ionizing solvent, water, by a non-polar organic solvent can result in improvements in the chemical purity of the metal values produced. Two methods of metal recovery from organic phases are available at present: hydrogen reduction and hydrolytic stripping, both methods being applicable to metal-loaded cationic extractants, in particular to carboxylic acids such as Versatic acid. Hydrogen reduction involves reaction at elevated temperatures (higher than 140 °C) between metal-loaded organic acids and hydrogen gas, yielding elemental metal:

R M + 2 M nRH, n (org) 2 (g) (s)

The pioneering work of Burkin on the reduction of nickel, cobalt, and copper from carboxylic acids and alkyl phosphoric acids is reviewed, together with more recent Canadian work on the recovery of copper powder from organic solutions containing the chelating extractant, Kelex 100. In some process situations, the recovery of elemental metal may not be necessary, and in these circumstances the recently developed hydrolytic stripping method may be applicable. This involves hydrolytic reaction at elevated temperatures between water and metal-loaded carboxylic acids, and yields metal oxides or hydroxides, e.g., for a trivalent metal,

+ 3H,O -» M,O,, + 6RH 2R,M(org) v (org)" Versatic acid is a suitable organic medium for these reactions, and metals which can be recovered by this method include iron, nickel, and copper, which precipitate as Fe2O3, Ni(OH)2, and CuO or Cu2O, respectively. In both types of reaction, one of the key advantages over conventional aqueous-phase precipitation is that there is no dissociation in the organic phase, and thus no free hydrogen ions are formed. Instead, the undissociated organic acid is regenerated, and there is no necessity to have a neutralizing agent present to consume hydrogen ions. The organic systems therefore contain fewer chemical components than the equivalent

86 aqueous systems, with the result that purer metal products can be obtained. This is \ of particular importance in the precipitation of hematite, which is the most interesting { potential application of hydrolytic stripping in hydrometallurgy. The great affinity of j carboxylic acid extractants for ferric iron allows very pure iron-loaded organic phases rj • to be produced by conventional solvent-extraction techniques. These are reacted with | pure water during hydrolytic stripping, and thus the hematite produced is ' ; uncontaminated by other inorganic impurities. In principle, this will enable further ' use to be made of the hematite in iron-making processes, or even in more specialized applications such as pigments. This is in contrast to the iron residues produced currently by conventional aqueous-phase hydrolysis of iron-bearing solutions, such as jarosite and goethite residues, where the products cannot be used and can give rise to severe , disposal problems. The kinetics of the hydrolytic stripping of iron were studied in detail. The precipitation of hematite can be initiated either by a process of homogeneous self-nucleation or by heterogeneous reaction at the surface of existing hematite particles. The former process is favoured at high loadings of iron in the organic phase and at high reaction temperatures. Both modes of precipitation exhibit high activation energies—91 and i 124kJ/mol for homogeneous and heterogeneous reactions, respectively—indicating | chemical-reaction control in both cases. The reaction rates are independent of the volume •-"~" ratio of water to organic phase, suggesting that the reactions involve water dissolved i in the organic phase and do not occur at the aqueous-organic interface. I The rate of the heterogeneous reaction is directly proportional to the surface area I of solid hematite present, and the kinetic data can be fitted by a Langmuir-Hinshelwood • model. The mechanistic interpretation involves a surface hydrolysis reaction between an iron-carboxylate complex adsorbed on one active surface site, and a water molecule adsorbed and dissociated on two sites. The change from heterogeneous to homogeneous self-nucleating behaviour with increasing temperature and increasing iron-loading in the organic phase is related to changes in the nature and distribution of organic iron- carboyxylate complexes. These changes were observed by the use of high-temperature infrared spectroscopy. If the organic phase contains more than one metal, it is possible to precipitate complex metal oxides by hydrolytic stripping for certain combinations of metals. Recent work on the production of magnetic ferrite oxides by this method is reviewed briefly.

87 Paper B5/2

) THE SEPARATION OF CHROMIUM AND IRON BY THE SOLVENT ! EXTRACTION OF FERROCHROMIUM LEACH LIQUORS

J.S. Preston Council for Mineral Technology South Africa The separation of chromium and iron by the solvent extraction of liquors obtained from the leaching of ferrochromium with either hydrochloric or sulphuric acid was investigated as part of a potential process for the production of electrolytic chromium metal from ferrochromium fines. The specification for a suitable electrolyte was set at a maximum of 100 p.p.m. of iron in a solution with a chromium content of 100 g/1. The leaching of ferrochromium fines with either hydrochloric or sulphuric acid produces solutions containing predominantly chromium(III) and iron(III). However, most of the tests were directed towards the selective extraction of iron(lll) on account of its relatively strong extraction by a variety of commercial reagents. Only a limited amount of work was conducted on the extraction of iron(II) since, although the selective removal of this species would eliminate an oxidation step, the generally much weaker complex- forming ability of ions of lower oxidation state renders the development of an iron(II)- selective system less probable. Also, only limited consideration was given to the selective extraction of chromium(IH) owing to its kinetic inertness towards the formation of complexes, and of chromium(II) on account of its facile oxidation by air. In sulphate media, iron(IH) was found to be separated from chromium(IIl) most readily by amine extractants. However, since the solvent loadings are relatively low, an initial partial separation of the components of the leach liquor is advantageous. This can be achieved by the crystallization of iron(IH) sulphate heptahydrate, by precipitation and re-dissolution of chromium(IH) hydroxide, or by a combination of these methods. After the solution has been oxidized, residual iron(III) can be extracted efficiently by, for example, primary amines:

(RNH3)2SO4 + 2 Fe(S%)2" = 2 RNH3Fe(SO4)2 + SO^.

However, as a result of the high formation constants of the anionic iron(HI) sulphate complexes, the equilibrium cannot be reversed by contacting of the loaded organic phase with water alone. The use of sulphuric acid as a stripping solution results in the displacement of the complex anion by the bisulphate ion:

RNH3Fe(SO4)2 + HSO4~ = RNH3HSO4 + Fe(SO4)2-

Secondary and tertiary amines were found to be more readily stripped than primary amines, but showed very poor phase-separation properties. However, the addition of suitable amounts of aliphatic alcohols to the primary amine solvents modified their extraction properties sufficiently to allow stripping with 1,0 M sulphuric acid.

88 Nevertheless, such a process suffers from the disadvantage that further treatment of the acidic iron solution would be required prior to its disposal. More promising results were obtained by the use of liquors resulting from the leaching of ferrochromium with hydrochloric acid. Extractants of the primary and secondary amine type showed high capacities for the extraction of iron(III) from concentrated chloride solutions, enabling such leach liquors (following oxidation with chlorine) to be treated without recourse to a prior separation step. In contrast to the situation in I I sulphate media, the equilibrium

RNH3C1 + FeCl4- = RNH3FeCl4 + Cl- ean be reversed at low aqueous concentrations of chloride, owing to the low degree of formation of the extractable FeCl^anion under such conditions. This enables iron to be stripped from loaded organic phases by water alone, and it was found that strip liquors with an iron concentration of more than 100 g/1 could be produced under suitable conditions. Such liquors might be suitable for conversion to hydrogen chloride by pyrohydrolysis (for recycling to the leach) and hematite (for sale or disposal). The paper presents results of small-scale tests on continuous countercurrent extraction from the liquors resulting from the leaching of ferrochromium fines with simulated spent electrolyte containing hydrochloric acid and chromium(III) chloride.

Paper B5/3

ZINC, MANGANESE, AND SULPHURIC ACID RECOVERY BY SOLVENT EXTRACTION FROM SPENT ELECTROLYTE

by

D. Buttinelli, C. Giavarini, and A. Mercanti University of Rome Italy

Previous papers have shown that tri-n-butylphosphate (TBP) and di-(2-ethylhexyl) phosphoric acid (DEHP) can successfully be employed for zinc recovery from acidic- waste liquors such as spent electrolytes of the zinc industry. A drawback to the use of TBP may be that sodium chloride must be added as a salting agent and that the extracted species are zinc chloride complexes. In the present work a different approach was tried, i.e. the preliminary extraction of sulphuric acid by iso-butyl alcohol (IBA) in order to reduce the acidity of the original liquor and to recover the acid itself, followed by the recovery of zinc by naphthenic acid (NA) or DEHP. The use of NA instead of DEHP is interesting because of the lower cost of the former and of the possibility of the co-extraction of manganese and zinc. Considering the low price of sulphuric acid, its recovery would probably not be of

89 J

interest unless we consider (1) the problem of its neutralization (before discharging the waste electrolyte), with a consequent consumption of chemicals, sludge formation, and metal losses; and (2) its re-use in zinc (and manganese) stripping from the organic solvent. The investigations were carried out first in batch, by using separatory funnels, and then in a small-scale continuous pilot plant, which consisted essentially of a battery of 16 mixer-settlers for sulphuric acid recovery and of 4 different mixers and settlers with intermediate pH control (plus 2 to 3 steps for final stripping) for zinc recovery (Figure 1). The flowrate of the two phases was 3 to 41/h. An industrial spent electrolyte was used containing the following: sulphuric acid 232 g/1, zinc 14,5 g/1, magnesium 17,8 g/1, manganese 1,4 g/1, and minor amounts of iron, copper, cadmium, nickel, arsenic, antimony, etc.

Iso-butyl alcohol

Mixer-settlers (8) Mixer-settlers (8)

Extraction of Stripping of H2SO4 I solution Spent H,O I(3 and 4 N electrolyte 1 Recovery of iso-butyl Recovery of alcohol Iso-butyl alcohol iso-butyl Calcined alrohol blende I O pH control Neutralization|—J Recovery of zinc solution To . electro*ysjs plant _^_ Solid residue A To Extraction of to treatment I waste" zinc (with pH control)

1 M naphthenic acid (or DEHP) in kerosene FIGURE 1. Flowsheet of the process tested

The pilot-plant tests showed that sulphuric acid could be extracted, leaving a minimum content in the electrolyte of 20 g/1, by using a high organic-aqueous (O/A) ratio and many steps; however, the corresponding acid concentration in the solvent was fairly low. A residual sulphuric acid concentration of 30 g/1, obtained with an O/A ratio of 10 and 7 to 8 steps, was considered to be more convenient: in that way an aqueous solution containing 150 to 170 g/1 of sulphuric acid was obtained after stripping with IBA; the solutions contained only traces of metals such as zinc, manganese, and magnesium. The electrolyte obtained after sulphuric acid recovery was partially neutralized to pH 4,5 with calcined blende to allow an efficient extraction by DEHP, or to between pH 5,2 and 5,5 when using NA. The use of blende instead of other neutralizing agents

90 was preferred in order to enrich the zinc content of the solution and to limit the formation of zinc precipitate. The extraction was performed with 1 M NA or DEHP in kerosene (O/A ratios of 1,5 to 2,5 being used): 3 to 4 extraction steps with intermediate pH control were necessary for NA, and only 2 for DEHP, to recover 96 to 99 per cent of the zinc; when using NA, 50 to 90 per cent of the manganese was also recovered, depending on the pH value. Solvent stripping was carried out with the recovered sulphuric acid solutions: one or two steps were sufficient to re-extract the zinc and manganese completely; by using an O/A of 0,5, a solution containing 30 to 40g/l of zinc and 1,0 to 2,0 g/1 of manganese was obtained finally. Apart from iron and cobalt, other impurities were not extracted. Solvent losses due to the solubility in the stripping solution were normally lower than 15 p.p.m. for NA and less for DEHP.

Paper B5/4 A REVIEW OF THE DEVELOPMENT OF RESINS FOR USE IN HYDROMETALLURGY

by

B.R. Green Council for Mineral Technology South Africa The use of anion-exchange resins for the extraction of uranium is well established. Similar resins have been investigated extensively for the recovery of gold cyanide and are employed for this purpose in some countries. In other areas of hydrometallurgy, the applications for ion exchange have been small and varied. It has happened that, where manufacturers anticipated large applications for resins, they were prompted to develop special resins. However, because of the volatile nature of the metals market, which may result in cuts in production and the abandonment of projects, their efforts were often in vain. Research and development in the area of resins for hydrometallurgy have therefore become unpopular, probably to the detriment of the industry. Because of the belief of the Council for Mineral Technology (Mintek) that special resins have a part to play in hydrometallurgy, it has continued to conduct work in this area. Mintek has considered a number of processes in which ion exchange or chelating resins can be used. Those which involve the recovery of either a particular metal or a group of metals are discussed, and resins (available or undergoing development) that may be suitable for the purpose are described. Because a variety of resin types and applications (involving different groups of metals) has been considered, the information that has been accumulated may be useful to scientists working in areas that have not been of direct interest to Mintek. Although ion exchange is well established in the processing of uranium and a number

91 'J

\ of strong-base resins are available specifically for the process, some attempts have been } made to develop special reagents to overcome certain problems. These include the high- ] density resins intended for use in columns that employ fluidized beds. The extraction 5 of uranium (possibly simultaneously with that of gold) by the use of weak-base resins ; has been of interest at Mintek, the aim being the preparation of resins with basicities : sufficiently low to allow elution to take place between pH 3 and 5. Interest in the use of resins for the extraction of gold from cyanide liquors has revived with the successful implementation of carbon-in-pulp processes. No resins have yet been developed specifically for this application, but there are indications that manufacturers are considering various aspects that may be important in resin-in-pulp processes such i as the durability of resin beads and the manufacture of large beads for easy screening. At Mintek, efforts have been directed towards the development of weak-base anion exchangers that will perform effectively at the pH value of the leach liquor, and that can be eluted easily with aqueous sodium hydroxide. Factors affecting the ability of a resin to be eluted have been investigated. The processing of low-grade oxidized ores containing base metals was probably the i first major area identified for the potential use of selective chelating resins. Although ! resins of this type with imminodiacetate groups had been available for some time, their usefulness was restricted because of their poor selectivity against iron. Efforts were made by a few researchers to develop reagents specifically for this type of application, and a number of interesting resins were prepared. The interest in chelating resins at Mintek includes their design and development, and the investigation of their properties and also those of commercially available resins. Resins have also been evaluated for possible use in specific applications. Because of the relatively small markets and high cost of these reagents, a new approach involving the use of a more versatile resin matrix already containing a nitrogen donor was investigated recently. A problem involved in the recovery of manganese (which is normally very weakly bound to most chelating resins) from the thickener underflow of an electrolytic manganese plant led to the development of a chelating resin that will load manganese at pH values as low as 3. This resin may also be useful for the extraction of cobalt, zinc, and iron(II) from fairly acidic solutions.

92 SESSION B6 Separation Processes (continued)

Paper B6/1

A RESIN-IN-LEACH PROCESS FOR THE EXTRACTION OF MANGANESE FROM AN OXIDE

by

M.W. Johns and A. Mehmet Council for Mineral Technology South Africa Because filtration is a major cost item in hydrometallurgical plants, considerable attention has been focused on processes in which this step is not required. This paper describes a resin-in-leach (RIL) process that was developed at the Council for Mineral Technology (Mintek) for the recovery of manganese by a weakly acidic chelating resin. The material from which the manganese is recovered is a sludge from the dust scrubbers above a ferromanganese furnace. In the resin-in-leach (RIL) circuit, free protons leach the ore, freeing manganese cations, which are absorbed onto the protonated resin, and protons are released that, in turn, are used for further leaching of the ore. The paper gives results that were obtained in laboratory tests on this process. In batch tests carried out at different resin-to-pulp ratios and pH values, the extraction of manganese from the feed was 90 per cent, up to 88 per cent of the extracted manganese being adsorbed onto the resin. A study was then made of leaching of the pulp in a sulphuric acid medium at a high temperature, and a total extraction of 98 per cent was obtained. After the leach, the sludge was neutralized since ion exchange is limited to acid concentrations (above a pH value of 1 to 2). Subsequent countercurrent contacting of the leached sludge with resin resulted in a total manganese recovery (over the whole circuit) of 97 per cent. The resin, which had been loaded with 35 g of manganese per litre, was eluted with sulphuric acid. Calcium had also become loaded on the resin, and it was found that it could be displaced by the contacting of the resin with a solution high in manganese. A conceptual flowsheet is presented for a plant producing 60001 of electrolytic manganese dioxide (EMD) per annum, together with estimates of inventories and the sizes of major items of equipment. The plant consists of a multi-stage countcrcurrent RIL adsorption section, a calcium-scrubbing section, a NIMCIX elution column using return electrolyte from an EMD cell as the eluant, and a stage for the purification of the electrolyte feed to an EMD cell.

it. J

Paper B6/2 ! STUDIES ON THE MECHANISM OF GOLD ADSORPTION ON CARBON

\ ' by

• N. Tsuchida, M. Ruane, and D.M. Muir Murdoch University Australia A basic study was made of the effect of cations, anions, oxygen, and organic solvents on the adsorption and stripping of Au(CN)2~ from activated carbon. The aim of the . study was to quantify and rationalize the importance of dissolved salts, free cyanide, and adsorbed oxygen on the loading of gold onto carbon and to understand the role of deionized water or organic solvents in assisting the desorption of Au(CN)2" from carbon in the Anglo-NIM or organic-elution procedures. The results of this study led to a clarification of the mechanism of gold adsorption, and a proposed structure of the surface oxide active site on carbon based upon its chemical and electrochemical ! characteristics. 1 _. Clean demineralized coconut-shell and Norit carbons were equilibrated over a range of pH values with pure synthetic solutions containing about 10 p.p.m. of KAu(CN)2, KAg(CN)2, Hg(CN)2, and various salts such as CaCl2, MgCl2, NaCl, KC1, KCN, and CuCN. Comparisons were also made with deoxygenated solutions and carbons which had been heated to high temperatures under vacuum to remove chemisorbed oxygen. It was found that between pH 4 and 10, much more Au(CN)2" was adsorbed on carbon than its counter cation, and this was accompanied by a release of OH" to + solution. The molecular ratio of Au(CN)2~ : K was about 2,5:1. In the absence of gold, there was very little adsorption of cations below pH 10 but there was a significant increase in cation adsorption above pH 10. In the absence of oxygen, much less gold, + silver, and copper loaded onto carbon and the molecular ratio of Au(CN)2~:K was close to 1:1. By contrast, the adsorption of Hg(CN)2 was unaffected by oxygen. Calcium loaded onto carbon at a lower pH value in the presence of CN" and oxygen. These results indicate that about 40 per cent of the gold is adsorbed as a neutral + n+ ion pair M ... Au(CN)2~ or M ... [Au(CN)2"]n, and that oxygen plays an important role in the adsorption of cyanide and gold onto carbon. Oxygen decomposes cyanide ion and Au(CN)2", leading to the precipitation of AuCN (equations 1 to 3):

+ H2O + 2e -• CNO" + 2OH~ (1)

Au(CN)2- + V2O2 + H2O + 2e -* AuCN + CNO" + 2OH~ (2)

CNO" + 2H2O -> HCO3- + NH3 (3)

The results with Hg(CN)2 illustrate the affinity of carbon for neutral species rather than charged cyanide complexes. Decomposition of cyanide is supported by evidence 2 of ammonia and by CaCO3 precipitation in the presence of Ca * (equation 3). It is consistent with this mechanism that the presence of additional salts like CaCL2 and

94 NaCl leads to a small enhancement of both Au(CN)2 arid cation adsorption, whilst the presence of additional cyanide leads to a small decrease in gold adsorption due to competition for active sites and oxygen. This, and other evidence, suggests that the surface oxide active sites consist of a number of functional groups including phenol, and/or hydroquinone, carboxyl, chromenol, and hydroperoxide. It follows from this model that, in order to desorb gold from the active sites, it is necessary to convert any AuCn to Au(CN)2~ and any phenol groups to phenoxide salts by presoaking with a solution of sodium cyanide and sodium hydroxide. However, desorption of Au(CN)2" occurs only when the ion pairs dissociate + + in hot deionized water, taking both Au(CN)2~ and Na or K into solution. Results on the effect of water quality or ionic strength using the Anglo-NIM procedure are used to illustrate this point. Ground waters which contain significant quantities of cations like Ca2+ and Mg2+, which form strong ion pairs, are shown to elute gold relatively slowly. Organic solvents inhibit the adsorption and enhance the desorption of gold through a different physicochemical effect. Measurements of the activities of ions in various aqueous organic solvents show that the organic solvent increases the activity of cyanide ion relative to that in water and significantly decreases the activity of the Au(CN)2~ complex ion. This effect is most pronounced using acetonitrile as solvent. Thus, gold is eluted by ion exchange rather than ion-pair dissociation, and hence is not so dependent on temperature or ionic strength. High concentrations of gold can be eluted from carbon in 1 to 2 bed volumes at 25 °C using 40 per cent (v/v) acetonitrile and water containing 10g/l sodium cyanide.

Paper B6/3

PRACTICAL CONSIDERATIONS OF GOLD RECOVERY BY CARBON FROM MAIN AND EFFLUENT STREAMS AT VAAL REEFS

by

W.A. Kretschmer, R.F. Dewhirst, RJ. Human, RJ. Kelly, and B. Strong Vaal Reefs Exploration and Mining Co. Ltd South Africa

Vaal Reefs Exploration and Mining Company Limited (V.R.E.M.) is a major mining company situated in the Klerksdorp area, producing gold, uranium, and sulphuric acid in its seven main processing plants. Attention has been given to recovering gold using the carbon-in-pulp (CIP) and carbon-in-solution processes. As well as the existing processes described here, several design options are currently under consideration. In summary, the following plants utilizing gold recovery with carbon are currently operated at V.R.E.M.

95 J

Regenerated Aqueous-solution Plants i Recovery of gold from the regenerated aqueous stream of the uranium solvent-extraction \ , circuit is practised at two Divisions. This constitutes a high gold tenor (70 g/t) and low- i volume throughput (300 m3 per month) operation, and at the East Division 22 kg of I gold per month is recovered at an efficiency of more than 95 per cent. j Previously the regenerated aqueous stream was closed-circuited, being pumped intermittently along with the gold recovered from the calcine-treatment process to the gold-plant leaching circuit. However, the recovery percentage was v jknown, and the presence of high organic loadings often caused coating on the cloths or candles of the filters following gold precipitation. The plant consists of six 2 m by 0,6 m diameter columns each packed with 100 kg of le Garbone G210AS 816 carbon. Solution direction is upflow through the columns. Gold tenors are monitored at each stage, and columns are taken off line when the carbon becomes fully loaded. At present this is despatched to the Rand Refinery for treatment at a loading of 20 kg of gold per ton.

. Primary-filtrate Plant I At West Division—and at a plant under construction at East Division—gold is recovered from the filtrate of the non-acid dewatering filters situated between the gold and uranium plants. This primary filtrate, which contains recoverable soluble gold lost frc n the gold plant, was previously discharged to pay dams, reclaimed, and used for monitoring purposes. The plant at West Division consists of six reclaimed Permutit fixed-bed ion-exchange columns each loaded with 4500 kg of G210AS 125 carbon. Run in two trains of three columns at design throughput, 240 000 m3 of solution per month at a tenor of 0,15g/m3 is fed in a down-flow mode through the plant. The recovery to date has exceeded 93 per cent, and about 15 kg of gold is produced per month. Carbon, at a gold loading of 5 kg/t, is despatched to the Rand Refinery for treatment. Feed solution to the columns must be clarified to avoid blinding of the carbon, and this renders barren solution suitable for applications such as gland-service water supply, thereby reducing fresh-water requirements.

Afrikander Leases Plant (A.F.L.) At A.F.L., a 50 000t/m main-stream CIP plant treats cyanided gold-bearing pulp after semi-autogenous milling. The plant currently treats ore at a head grade of 3,5 g/t, and recovers 94 per cent of the gold. Five stages of CIP contact vessels with a total inventory of 401 of G210AS 816 carbon recover the gold, which is then eluted and electrowon onto stainless-steel cathodes in Mintek-designed cells. < Major commissioning difficulties were initially experienced with various plant screens, ; botn pre- and post-adsorption, coupled with a very high proportion of extraneous wood i fibre and plastic material arising from the mining process. Changes to these units and circuit modifications have now successfully resolved these problems, and the plant is treating the design throughput.

96 Heap Leaching (South Lease Area) ] A small pilot plant has been established at the South Gold Plant to treat low-grade J (1,2 g/t) 'waste washing fines, and recover the gold on carbon following heap leaching. i :, The plant consists of an impervious pad of 11 m by llm onto which a heap of 2501 \ of material is built; 2,5 m3 of cyanide solution (0,035 per cent) per hour at a pH value j of 11,5 is pumped onto the heap and distributed via sprays. The pregnant solution thus : \ generated is pumped to a 350 mm by 2 m conical-bottomed carbon column, whilst the j overflow is recycled back to the heap. In an initial test run, a gold recovery of 52,4 per cent was recorded. A preliminary capital estimate for a full-scale plant to treat the total arisings of 20 000 t/m is R600 000, with working costs of approximately R2 per ton

Paper B6/4

A NOVEL APPLICATION OF RESIN-IN-PULP IN THE METALLURGICAL INDUSTRY L. by C.A. Fleming and G. Cromberge Council for Mineral Technology South Africa

( A number of dump-reclamation projects are being undertaken in South Africa. The procedure that is generally adopted in the treatment of these residues is flotation of a pyrite concentrate and treatment of the concentrate for the recovery of gold, uranium, and sulphuric acid. The tailings from flotation, which can contain up to 50 per cent of the gold and 80 per cent of the uranium, are generally pumped to waste without further treatment. In certain circumstances, it has become economically viable to treat these flotation tailings, and one of the most promising flowsheets, from the standpoint of both capital and operating costs, is a combined resin-in-pulp (RIP) process in which gold and uranium are extracted from the tailings by an anion-exchange resin. Both gold and uranium are leached as anionic complexes (Au(CN)2~ and UO2(SO4)*" respectively), and both metal complexes can therefore be extracted by the same strong- or weak-base resin in a single RIP plant, thereby dispensing with two filter plants. Moreover, techniques have been developed that permit the two metal complexes to be < eluted from the loaded resin, either together or separately, to produce concentrated • gold and uranium streams for further processing. ' A number of configurations and flowsheets are possible in the combined RIP process. (1) The aurocyanide anion is fairly stable under the acidic, oxidizing conditions required for the dissolution of uranium. Therefore, it should be possible for gold to be leached in alkaline cyanide solution, and then for the uranium to be leached

97 after the pH value of the pulp has been reduced to around 2, while both me*al ! complexes are kept in solution in an anionic form. Then, after the double-leaching \ operation, both metals could be extracted in a single RIP plant. ! (2) Alternatively, the gold cyanide could be extracted from the alkaline-leach pulp I by resin prior to acidification of the pulp for uranium leaching. The separation ( of the leaching and RIP for gold from the operations for uranium would increase the number of RIP absorption contactors, but would allow greater flexibility in that each of the two RIP plants could be optimized to meet the specific demands of the particular metal complex. For example, the concentration of uranium would generally be about a hundred times greater than that of gold, and the flowrate ; of resin would need to be greater in the uranium RIP plant than in the gold plant for the same overall extraction efficiency. Another factor that would need to be taken into consideration is the effect of each metal complex on the resin-loading characteristics of the other, and these factors would be difficult to optimize in a combined extraction process. , (3) Another alternative, which would be viable where the leaching kinetics of gold | and uranium are fast, would be a combination of the leaching and extraction operations in a resin-in-leach application. This would minimize the number of adsorption contactors, as exemplified in the flowsheet for (1) above, while allowing for the maximum flexibility, as seen in the flowsheet for (2). Three different feed materials were tested in a small-scale (0,11 per day) RIP pilot plant in an evaluation of the various flowsheets, and the feasibility of the process with respect to the chemistry of leaching, extraction, and elution was demonstrated. In the first campaign, in which flotation tailings from a dump were treated, the concentrations of gold and uranium in solution after leaching were 0,15 and 30 p.p.m. respectively. The feed material in the second campaign was a concentrate resulting from the treatment of an old dump by wet high-intensity magnetic separation, and the concentrations of teachable gold and uranium were found to be 1,1 and 500p.p.m., respectively. Finally, the normal underground material being treated on one of South Africa's gold-and- uranium mines was examined, the concentrations of gold and uranium in solution after leaching being 3,3 and 300p.p.m., respectively. The paper gives results from the pilot-plant campaigns, and discusses the various flowsheets in relation to such factors as overall recovery of metal, consumption of reagents, and simplicity of operation. In addition, results are presented for the elution of gold, uranium, and other metal complexes from loaded strong- and weak-base resins.

98 SESSION B7 Process Development

Paper B7/1

RECENT ADVANCES IN THE DEVELOPMENT OF HYDROMETALLURGICAL PROCESSES FOR THE TREATMENT OF BASE METAL SULPHIDES

by

E.D. Nogueira Technicas Reunidas, S.A. Spain This paper deals with the hydrometallurgical treatment of base-metal sulphides (specifically copper, zinc, and lead) from materials in which precious metals, mainly silver and gold, may be present in significant amounts. Most of the research-and- development effort devoted to this matter in the past twenty years has been confined to two broad types of processes that are characterized by the kind of oxidizing agent used to leach the metal sulphide, namely, chlorine and oxygen. Within the first category are the very well-known processes of Clear-Duval, Elkem, Minemet, Bureau of Mines, Dextec, etc. All of these use a chloride medium for the leaching, purification, and recovery of the metal values, taking advantage of the different features offered by chloride chemistry in the achievement of these goals. The second type of process includes ammoniacal leaching of sulphides as in the Arbiter process, pressure leaching with air at low temperature as in the Sherrit-Cominco process, and pressure leaching with oxygen at high temperature as in the Comprex process of Technicas Reunidas. In all these processes, the metal sulphides are converted into metal sulphates. The paper introduces some recent advances in this field. The first group relates to pressure leaching. Several new applications of this technique, such as upgrading of copper concentrates, treatment of copper-lead concentrates, and recovery of copper and silver from a tetrahedrite concentrate, are presented together with some experimental conditions and results. In addition, a process for the treatment of bulk flotation concentrates from complex sulphide ores having a relatively high zinc content (20 to 30 per cent) is described in detail. This process is based upon the merging of pressure- leaching techniques at low and high temperatures, and is a good example of recent advances in the leaching of complex -sulphides. The second group of advances concerns the selective solvent extraction of metal chlorides, combined with a new electrowinning cell for the simultaneous production of base metals and chlorine. This offers a new approach to process developers in chloride hydrometallurgy for the production of base metals from sulphide concentrates, since free chlorine becomes available for sulphide leaching. Selective solvent extraction of metal chlorides made a great advance recently through the development of new extractants, such as various phosphonates for the extraction of zinc chloride, and DS-5443 (by ICI) for the extraction of cupric chloride. Basic data are reported for both systems. The new electrowinning cell, known as the METCLOR

99 cell, utilizes a cation-selective ion-exchange membrane to separate the anolyte from j the catholyte. It can be used in principle with any metal chloride solution providing ; that a suitable selective extractant is available to achieve the required degree of solution • purity to feed the cell. i Finally, several processes for the production of zinc, lead, and copper are described I in detail, including some experimental results. These are presented as examples of process development in hydrometallurgy based upon recent advances in solvent extraction and electrowinning.

Paper B7/2

THE CARBON-IN-PULP PLANT AT RAND MINES MILLING AND MINING COMPANY LIMITED: PROBLEMS ENCOUNTERED AND DEVELOPMENTS INTRODUCED

by

j P.A. Laxen* and T.D. Brownf * Council for Mineral Technology tRand Mines Milling and Mining Co. Ltd South Africa

When the gold-recovery plant of Rand Mines Milling and Mining Co. Ltd (RM3), situated on the Crown Mines property near Johannesburg, was commissioned in September 1982, the carbon-in-pulp (CIP) circuit represented the largest such installation in South Africa and was, in fact, one of the largest in the world. This plant had been designed to treat reclaimed sand and slime from dumps sited mainly on the old Crown Mine. The layout of the plant is described. The plant had to contend with a number of unusual features. Firsdy, the large tonnage to be handled in the CIP circuit, approximately 380 0001 of solids per month, necessitated the development and introduction 'on site' of a number of alterations in the interstage screening of pulp and the movement of carbon. These developments, which are described and discussed, could well be applicable in future CIP plants. Secondly, the oxidation of sulphides in the duimps had produced soluble salts. The flowsheet was designed to reject these salts, but a small proportion entered the cyanidation circuit and affected cyanidation. Under certain circumstances, these salts were loaded onto the carbon in the adsorption circuit. The solutions to these problems i are discussed. ; Thirdly, some of the water used in the RM3 plant was river water, which was found t to contain varying quantities of organic materials that caused a number of problems in the adsorption and elution stages of the CIP circuit. The possible presence of organic matter derived from the material in the dumps was also investigated. Fourthly, it was unexpectedly found that all the eluted carbon on each cycle had to be regenerated, and the regeneration kiln did not have the necessary capacity. This

100 aspect, which led to the recycling of poorly regenerated carbon, and hence contributed to the poor adsorption of gold, is discussed. A new reactivation furnace of novel design was introduced into the RM3 circuit. Fifthly, a sulphide-flotation step in the RM3 circuit required that the flotation concentrate and flotation tailings should be cyanided separately and treated by CIP in separate adsorption circuits. The ramifications of this flotation step are discussed. Sixthly, severe problems occurred at one stage with the filtration of the zinc precipitate. (The batch eluate at RM3 was treated for the recovery of gold by precipitation with zinc, followed by filtration, calcining, and smelting.) The cause of these problems was ascertained, and improvements were made to the elution procedure. A technique of split elution for the recycling of the caustic-cyanide reagents was introduced, being tested first on pilot-plant scale in a laboratory. From the start of the operations, a pilot plant was operated at RM3 alongside the main CIP plant. This pilot plant proved to be most useful in determinations of the nature of the problems being encountered and of ways in which they could be solved. The results generated during the operation of the pilot plant are presented and discussed. By the concentration of a great deal of effort on the CIP plant at RM3, an improved understanding of the variety of problems likely to be encountered in such an operation was obtained, and, as the final result, the CIP operation was optimized.

Paper B7/3 THE LODEVE MILL: A COMPLEX ALKALINE PROCESS FOR A COMPLEX URANIUM ORE

by

R. Bodu*. J.P. Herbert*, P. Lafforgue*, G. Lyaudet*, P. Michelf, and R. Vanhelleputtet *COGEMA fSIMO France Geology and Mineralogy Uranium mineralization occurs in a Permian formation including shales and carbonates. They can be reduced to two types: stratiform and located in layers rich in organic matter, or in veinstones close to faults, associated with asphaltic matter. The Lodeve ore is difficult to process due, for a large part, to the organic matter, which is often associated with uranium, and to the various constituents of the ores, which are as follows. Natural carbonates: mainly enkerite (ferriferous magnesian limestone) and siderite. Owing to their presence, the process uses alkaline leaching. Phyllite materials (such as montmorillonite) Sulphides: pyrite probably the main constituent, with other sulphides (such as blende), sulphates, and some sulphur bound to the organic matter

101 Molybdenum: bound to the uranium mineralization only in part and can appear as ilsemanite. Zirconium: seems to be bound to complex silicates rich in uranium.

Process Studies Investigations started in 1975, when drill cores became available. On a laboratory scale, tests were carried out in autoclaves of 1 to 2 litres effective capacity on ore samples of 1 kg. The main features of the process were so defined. On a pilot scale, a large autoclave (effective capacity 750 litres) fitted with a twin-propeller mixer was used. Solid-liquid separations were carried out on two 0,5 m2 belt filters. To define the process, the following aims were fixed: Uranium dissolution efficiency as high as possible Production of a yellow cake meeting the refining specifications Molybdenum recovery as sulphide Processing of solid and liquid wastes as required by French environmental regulations: salt content, pH, suspended particles, uranium, and radium. To achieve these aims, the following process steps were defined: Grinding: rf95 = 160 pm Two stages of alkaline leaching under a pressure of 6 atm and a temperature of 140 °C, pure oxygen being used for uranium oxidation Solid-liquid separation on belt filters Precipitation of calcium uranate and calcining Dissolution of calcium uranate and precipitation of a final concentrate (magnesium uranate) Crystallization of sodium sulphate from the main liquid waste.

Problems in Design and Construction of the Lodeve Plant The main problems were encountered in the leaching of ore, reduction of the volume of solutions, and processing of the effluent. Dissolution of uranium: The equipment necessary to the following operations was as follows: Grinding and classifying system achieving the right size distribution Leaching carried out in 18 vertical autoclaves (45 m3 each) Solid-liquid separation after each leaching stage carried out on 6 belt filters (40 m2 each). The recycling of solutions and use of appropriate pumps to restrict the volume of solutions produced. Saving of energy: A new type of heat exchangers was used, resulting in the recovery of heat from slurries coming from the autoclaves. Vapour from the evaporation of effluent is compressed mechanically and recycled. Processing of uranium solution: The process flowsheet for uranium solutions is more conventional. However, a multiple-hearth calciner is needed to burn the carbonaceous matter contained in the preconcentrate.

102 Processing of effluent: The main effluent, rich in sodium sulphate, must be evaporated and the sulphate crystallized, resulting in a production of 25 kt per year. To be sold, the sulphate must meet very strict specifications. Several units have therefore been designed and built to purify the sulphate before being crystallized. Other problems: Most of the other problems were related to the environmental regulations and to the short time allowed for the erection of the plant.

Industrial Operation Starting up a plant in which a conventional process is applied usually needs some time. With a new process, that time is longer. At the Lodeve plant the achievement of the designed throughput has taken almost two years, and many improvements are being made at present. The main problems have been met as follows: Grinding unit: The right size distribution has not yet been achieved; the classification loop is now being improved. Leaching unit: It was started up carefully but did not cause many difficulties. Solid-liquid separation: The throughput achieved at pilot scale has not yet been obtained at full scale. The problem seems to be due to an excess of fine particles in the slurry (see grinding). I Solution-processing unit: Most problems have appeared with and around the calciner. As the yellow cake produced does not meet the specifications, the process will be largely modified. Liquid effluent: Because each operation depends on the previous one, any difficulty in one part of the process can disturb the whole unit. Most of the difficulties have now been overcome thanks to the collaboration of the operators at COGEMA and SIMO.

Paper B7/4

THE PRODUCTION OF ELECTROLYTIC MANGANESE DIOXIDE FROM FERROMANGANESE FURNACE SLUDGE

by

C.F.B. Coetzee and W.A.M. te Riele Council for Mineral Technology South Africa The Republic of South Africa has large deposits of manganese ore and relatively cheap electric power. The local production of battery-grade manganese dioxide therefore appears to have natural advantages. There is a growth in the demand for dry-cell batteries of 4 to 5 per cent per annum. Naturally occurring ores of battery-grade manganese dioxide and chemically produced

103 ) manganese dioxide, although less expensive to produce than electrolytic manganese J dioxide (EMD), are not favoured for 'heavy drainage' dry-cell applications. EMD ) performs well in such batteries, and its particle size, which is an important factor in \ the manufacture of dry-cell batteries, is simpler to control. j The minerals normally used as starting materials for the production of EMD are : rhodocrosite and pyrolusite. However, during the production of ferromanganese, the furnace gases carry some of the reduced manganese fines (predominantly in the form of acid-soluble manganous oxide) to the scrubbers. This sludge is a potential source of acid-soluble manganese at zero cost. In a study conducted at the Council for Mineral Technology (Mintek) into the use : of furnace sludge for the production of EMD, the following basic process steps, which are similar to those reported in the literature, were investigated: acid leaching to produce a manganese sulphate liquor, solid-liquid separation, and purification and electrolysis. The characteristics of the sludges were determined with regard to teachability, acid- consuming constituents, impurities, particle size, filtration, etc. The objectives of the study were as follows: ! to prove that battery-grade EMD can be produced from furnace sludge, L to compare the production costs of EMD from furnace sludge with those of manganous oxide reduced from pyrolusite, and to produce sufficient EMD for submission to battery manufacturers for evaluation. The experimental work covered various process routes. Single-stage leaches were followed by the use of lime or excess sludge for the neutralization of the leach liquor. A two-stage leaching procedure was tested in which the free acid in the first-stage liquor was neutralized with excess sludge or lime. Manganese extractions of 80 to 85 per cent were obtained. The consumption of sulphuric acid varied according to the method of neutralization. It was found that thorough washing of the filter cakes could reduce the dissolved losses to less than 2 per cent. Although the presence of alkali metals helped to remove most of the iron as jarosite, the filtration was poor owing to the fineness of the sludge particles. " A small electrolytic cell was operated to provide information on the anodes to be used, the acid levels in the cell, the materials of construction, etc. This unit operated at production rates of about 180 g of EMD per 24 hours under automatic control. The information gained from this plant was used to guide the operation of a pilot plant at a production rate of around 10 to 25 kg of EMD per day. Various problems were revealed that had not been encountered in the bench-scale work: these involved frothing during leaching, snapping of graphite electrodes, quality of commercial rods, contamination of electrolyte, etc. The manganese dioxide in the EMD produced initially was below the value accepted by industry, which is a minimum of 90 per cent, and some of the impurities from the electrolyte reported in the EMD. Two more pilot-plant campaigns were run using high-purity graphite rods and two sources of raw material: prereduced manganese ore and furnace sludges. The use of prereduced ore was found to have some advantages over the sludge in that there is no frothing, the acid consumption is lower, and the filtration is better.

104 Although the high-quality graphite anodes also improved the results, the manganese dioxide content of the EMD was still only just 90 per cent. 1 A further series of tests was run on the small cell to show the effect, on the quality of the EMD, of different acid concentrations in the electrolyte, of the use of titanium anodes varying in shape and surface treatment, and of organic material (coal-tar, etc.) in the electrolyte. Corrugated titanium anodes worked well, as did flat sheets with 5 mm holes drilled at intervals. The adhesion to flat sheets (without holes) was very poor, and the treatment of the titanium surfaces with sandblasting and a coating of ruthenium oxide made the anodes less prone to passivation. In the final pilot-plant run, the concentration of acid in the electrolyte was about 40 g/1, and activated carbon was used to remove any dissolved organic material from the fresh feed to the cell. High-grade EMD was produced at current efficiencies of 85 per cent.

105 SESSION B8 } Process Development (continued)

1 Paper B8/1 r j RECOVERY OF LEAD FROM MIXED SULPHIDE CONCENTRATES

by

A.O. Filmer and G.G. Briggs CRA Services Ltd ; Australia Complex sulphide ore-bodies are often not amenable to selective flotation to produce individual concentrates of copper, lead, and zinc. Values which are misplaced into another concentrate do not attract full payment from a toll processer, and often a penalty is incurred if their levels exceed prescribed limits. This paper presents a i hydrometallurgical process (Figure 1) for selective removal of excessive lead from copper i and zinc concentrates. This yields an upgraded feed to an existing process and lead ~~~ metal which is suitable for further refining. The process consists of a mildly oxidizing leach in acidic chloride solution to yield selective dissolution of galena:

+ PbS + 2H + 4CL" -> PbClJ- + H2S (1)

3+ 2+ + 2Fe + H2S -* 2Fe + S + 2H (2)

Chalcopyrite and sphalerite are not attacked under these conditions. The clarified pregnant liquor containing lead as the chloride complex is transferred to the cathode compartment of an electrowinning cell, where lead is plated as a powder:

PbClJ" + 2e~ -» Pb + 4C1" (3)

The lead powder is collected from the base of the cell. The barren catholyte is transferred to the anode compartment of the electrowinning cell, where iron(II) is oxidized to iron(IH):

2Fe2+ -• 2Fe3+ + 2e~ (4)

The solution is then returned to the leach to yield an overall reaction

PbS -» Pb + S (5)

A semi-continuous laboratory-scale reactor was run to illustrate the concept. The leach was controlled by adjusting the Ej, of the reaction to obtain selective dissolution. In the laboratory, this was achieved by adjusting the current to the electrowinning cell

106 Leaded concentrate

Leaching Liquid

Liquid Fe(II) oxidation Solid-liquid Pb recovery separation

Solid H,O 1 I Wash Pb T Deleaded concentrate

FIGURE 1. A flowsheet for the deleading process for constant rate of batchwise addition and removal of concentrate. It is envisaged, however, that the process could be fully continuous and controlled by the feed rate of concentrate. At 95 °C, with 250 g/I of sodium chloride, and an Eh of 30 to 50 mV (S.C.E.), lead extractions of 75 to 90 per cent could be achieved with a mean solids residence time of one hour. Less than 1 per cent copper and zinc dissolution occurred under these conditions. The lead powder produced at cathodic current densities of up to 500 A/m2 was of purity greater than 98 per cent. Consumptions of about 0,7 kg of hydrochloric acid per ton of concentrate and power requirements for the electrowinning of 0,6 kW*h per kilogram of lead were recorded.

107 Paper B8/2

| AN ALTERNATIVE ROUTE FOR THE PRODUCTION OF CHROMIUM { CHEMICALS FROM CHROMITE

| M.F. Dawson and R.I. Edwards Council for Mineral Technology South Africa

By far the largest end-usage of chromium is in the form of various ferrous alloys. Nevertheless, its consumption in the form of chemicals represents an important market in terms of quantity, and even more so in terms of value. A wide variety of chemicals is marketed, the most important product being sodium dichromate, but chromic oxide, chromic acid anhydride, and basic chromium sulphate are also produced in large tonnages. These last three chemicals are all derived from sodium dichromate, which is in turn manufactured from chromite by a well-established process universally used by the industry. In this process, chromite is roasted with sodium carbonate under oxidizing conditions at a temperature of 1100 °C. Sodium chromate is formed, and this is recovered by leaching of the calcine with water. The leach liquor is purified and acidified to produce a solution containing sodium sulphate and sodium dichromate. Sodium sulphate is removed by crystallization before concentration and crystallization of the product, sodium dichromate dihydrate. Although the process is simple in principle, it has a number of disadvantages in practice. These include a relatively low yield of chromium, high operating costs, the generation of large quantities of toxic effluent, and severe industrial hygiene problems associated with the handling in the dry state of hexavalent chromium compounds. These disadvantages motivated an investigation into alternative methods for the production of chromium chemicals by a more economic route better suited to South African conditions. Of the many possible routes that were investigated initially, a process consisting of reduction roasting and leaching with sulphuric acid appeared most promising and was selected for further investigation. In this proress, chromite is reduced in the solid state at relatively low temperatures (1200 to 1300 °C) to form finely divided ferrochromium carbide. This is leached in dilute sulphuric acid to produce a leach liquor containing Fe(II) and Cr(HI) at a pH value of about 2. Neutralization of this liquor with sponge iron precipitates chromium hydroxide in a granular form coarse enough to be filtered easily. The filtrate from the neutralization contains ferrous sulphate, from which hematite and sulphuric acid are recovered by high-pressure oxidation and hydrolysis, as used in electrolytic zinc plants. Sulphuric acid is recycled to the lea-rh, while the hematite byproduct is reduced to sponge iron for recycling to the neutralization step.

108 The chromium hydroxide product from the neutralization step has the following analysis: Cr 25 to 30 per cent, SO4 15 to 20 per cent, H2O about 50 per cent, and Fe less than 0,5 per cent. Sulphate can be removed by washing with dilute alkali to levels of less than 0,5 per cent. The product can be dissolved in sulphuric acid to produce basic chromium sulphate for sale, or for electrolytic conversion to chromium metal. Calcination at 1000 °C gives a product suitable for high-quality refractories and also for the production of chromium carbides and metal. It can also be converted into hexavalent chromium components (chromic acid, etc.) by anodic oxidation. The advantages of this route over the conventional one are as follows: (a) the consumption of reagents is lower (only char and chromite being consumed), (b) the process effluents are inert solids, posing no pollution threat, and (c) there are no health hazards associated with hexavalent chromium compounds. The process is at present being studied on a miniplant scale, but it is hoped that this will be followed by a pilot-plant campaign at a production rate of 50 kg of chromium hydroxide per day.

109 SESSION Cl Modern Steel Processes \ i Paper Cl/1 • ECONOMIC AND TECHNICAL ASPECTS OF THE DIRECT INJECTION OF i COAL IN IRON- AND STEEL-MAKING OPERATIONS I by L. Von Bogdandy, M. Chitil, J.A. Innes, and C.G. Jonker Kloeckner CRA, West Germany

Paper Cl/2 THE KR PROCESS, A NEW DEVELOPMENT IN STEEL MAKING

by

I E. Golde Korf Engineering GMBH West Germany The KR Process is a coal-reduction process for the production of hot metal and reducing gas, and can be based on a great variety of run-of-mine coals. The reduction of lump ore and pellets with natural gas as a reductant has been solved on a large industrial scale (Midrex) and today belongs to the state-of-the-art. As a result, the still dominating blast furnace has met with its first competitor after approximately two-hundred years. However, two problems remain to be solved in modern reduction technology: (1) the direct reduction of lump ore or pellets by the use of coal, and (2) the direct reduction of ore fines by the use of coal or gas. During the last century, many attempts were made to solve these problems. Rotary- kiln processes were considered a possibility for quite some time, and the combination of coal gasification and gaseous direct reduction also appears to be technically feasible. However, up to now the requirements of plant availability, low investment, and low production costs have not always been met by the combination of these processes. The KR Process, which has been under development by the Korf Group since 1978, aims at expanding the energy basis for integrated steel mills to low-quality non-coking coal, of which there is great abundance. Up to now, coking coal and natural gas provided the energy for the reduction of iron ore. Several test campaigns have been conducted in a pilot plant at Kehl on the Rhine with a capacity of between 5 and 101 of liquid iron per hour. Since 1979, the further development of this process has been a joint venture between the Korf Group and the Austrian company Voest-Alpine. The start-up took place in 1981-1982. Expressed in simplified terms, the KR plant represents a blast furnace consisting of a melter-gasifier and a reduction shaft furnace.

110 Paper Cl/3

THE BEHAVIOUR OF SULPHUR IN A KILN FOR THE PRODUCTION OF SPONGE IRON

by

J.A. Theron* and R.J. Dippenaarf •Council for Mineral Technology tUniversity of Pretoria South Africa The thermodynamics and kinetics governing the transfer of sulphur between coal, iron ore, and dolomite in a rotary kiln were studied. Research into the direct reduction of iron was motivated by Iscor's contract with Lurgi in which Lurgi undertook to erect a four-kiln sponge-iron plant at Iscor's Vanderbijlpark works. The rated capacity of the plant is 600 kt per annum, and production is due to start in 1984. As special steels are to be produced from the sponge iron, it is important that the sulphur content should be lower than 0,010 per cent. The study described in this paper started with a literature survey. The subsequent experimental work consisted of interrupted and simplified tests. In the interrupted tests, an attempt was made to simulate the transfer of sulphur in the kiln by the heating of mixtures of raw materials in a stainless-steel retort. Efforts were concentrated as follows: (1) during preheating of the charge, the heating rate, particle size of the iron ore, and charging of the dolomite in either the raw or the calcined state were varied, and mass balances for sulphur were compiled; and (2) the technique used for the desulphurization of sponge iron in the kiln by the addition of coal at the discharge end of the kiln was modelled by the fast heating in the retort of kiln-reduced sponge iron, coal, and, in certain cases, dolomite. In a simplified test, thermogravimetric apparatus was used in the study of a specific solid-gas reaction at high temperature. The design of the apparatus is such that an airtight box is positioned beneath a vertical-tube furnace. A crucible in the centre of the furnace rests on the balance weighing system with all friction in the arrangement carefully avoided. Gas mixtures containing nitrogen, carbon dioxide, hydrogen, hydrogen sulphide, carbon monoxide, and carbon oxysulphide can be prepared with a great deal of flexibility. Some of the results are of fundamental significance. Experimental work done with the thermogravimetric apparatus included the following: (i) the reduction of Sishen iron ore by hydrogen, (ii) the pick-up of sulphur and the desulphurization of sponge-iron platelets by the use of mixtures of hydrogen and hydrogen sulphide, (iii) the pick-up of sulphur and the desulphurization of hydrogen-reduced sponge iron and kiln-reduced sponge iron by the use of mixtures of hydrogen and hydrogen sulphide, (iv) the reduction of Sishen iron ore by carbon monoxide, and

111 (v) the pick-up of sulphur and the desulphurization of sponge iron by the use of mixtures of carbon monoxide and carbon oxysulphide. The experimental temperatures varied between 650 and 1050 °C. Between 900 and 1000 °C, the iron sulphide phase is molten and penetrates the porous structure of the sponge iron. For sponge-iron platelets, the desulphurization curve for 1000 °C levels off after approximately 40 to 60 per cent of the sulphur has been removed. At 900 °C, the deposition of sulphur is limited to the edge of the specimen, and the desulphurization is quantitative. Differences in the microstructure are being investigated as possible causes of this phenomenon.

Paper Cl/4

Title and abstract not yet available

112 SESSION C2 Furnace Development

Paper C2/1

THE OPERATION, CONTROL, AND DESIGN OF SUBMERGED-ARC FERRO-ALLOY FURNACES

by

M.S. Rennie Council for Mineral Technology South Africa For many years, the Council for Mineral Technology (Mintek) has been actively engaged in research-and-development work on submerged-arc furnaces. During this period, the South African ferro-alloy industry has grown considerably because of the natural abundance of chromium, manganese, and iron ores, and, initially, the availability of cheap electric power. Larger furnace units were constructed to take advantage of the economics of scale-up and the growing requirements of the World market. These large furnaces gave rise to problems associated with their size such as a lower power factor, greater interaction between phases, the importance of furnace permeability, and the need for careful electrode management. Furnaces were redesigned, operating techniques were improved, better measurements were introduced, and much tighter control was maintained over all aspects of furnace operation. Mintek was able to develop, assist with, or simply observe many of the improvements that were implemented over this period. Data accumulated from many furnaces and processes under a variety of economic conditions showed that there are common factors in the operation, control, and design of all these furnaces. These factors are discussed and applied to the production of four different ferro-alloys in hypothetical submerged-arc furnaces. The furnaces considered are those for the production of ferromanganese, ferrochromium, ferromanganese silicide, and ferrosilicon, and range in size from 20 to 40MW with electrode diameters of 1,3m, 1,5m, and 1,7m. The current-carrying capacity of a Sdderberg electrode depends on its diameter, which influences both the resistance and the power developed within a furnace. Under normal operating conditions, the passage of current is well defined, being between the electrode dp and the rnetal bath. The differences between processes are determined by the resistivity of the medium and the extent of arcing for that process. The resistivity of carbon-enriched slag, calculated with an on-line data-acquisition system by the use of the electrode as a probe and of its measured responses, was found to be very similar for a number of different ferro-alloy processes. This remarkable result allowed the calculation of the operating characteristics of these processes subject to the limitations on resistance that are obtainable in practice. These limitations depend on arcing, the carbon balance, and the distance between the electrode and the metal bath.

113 \ These are the very factors that must be controlled in the operation of submerged-arc i furnaces. f j . Based on the restrictions on the distribution of power within a furnace and the * '. I limitations on the electrode resistance, calculations of such factors as inter-electrode ^ • spacing, and the, internal diameter and depth of the: furnace, can be made.

Paper C2/2

OPERATIONAL PARAMETERS FOR SODERBERG ELECTRODES FROM CALCULATIONS, MEASUREMENTS AND PLANT EXPERIENCE

by R. Innvaer, L. Olsen, and A. Vatland Elkem R&D Centre Norway

lniroaucuon S^derberg self-baking electrodes for electric reduction furnaces have been in operation for more than sixty years. During this period, the electrode system has been increased steadily, in dimension and in current load. Today the largest electrodes are 2 m in diameter, weigh about 601, and conduct currents of more than 150 kA. Such electrodes demand comprehensive improvements in design as well as in operation. In addition, the electrode materials are exposed to very rigorous conditions in the furnace. If the electrode is unable to meet these requirements, problems will arise. The most serious, breakage of the electrode, will lead to unfortunate economic consequences for the whole furnace operation.

Objective The target of the research-and-development work reviewed in this paper was to secure a safe electrode operation. Factors like electrode current load, slipping rate, and shut down-start up procedures are discussed, together with the way in which they are related to the electrode diameter, equipment, material properties, and whether the furnace is open or closed. The results presented were obtained by use of a combination of mathematical models, measurements, and records on electrodes in normal operation._ The work was focused on an electrode of 1,7 m diameter in an open furnace for the production of ferrosilicon.

Position of Baking Zone In the holder area, electrode paste is baked to a material of high strength. The conditions are very important to heat generation and current distribution. A baking zone that is too low may lead to the most severe kind of breakage in the electrode, which would cause leakage of the soft paste. By the use of temperature measurements, the baking area was studied under different

114 electrode conditions, the results being used for the verification of temperatures computed in mathematical electrode programmes. Table 1 demonstrates the calculated effect on electrode baking of changes in the standard operating conditions. Only one parameter was changed at a time. The values show that the position of the baking zone—and the possibility of compensation for electrode consumption—is closely related to the electrode current and slipping rate. But other factors can also to some extent regulate the heat for electrode baking when changed within reasonable limits. TABLE 1

Calculations of baking-zone position in an electrode of 1,7 m diameter in a ferrosilicon furnace

Baking-zone position* Electrode conditions Standard Changed to mm Current 130 kA HOkA 215 Slipping rate per 24 h 0,6 m 1,0m 190 Steel thickness casing 2,7 mm 3,0 mm 250 Width casing fins 340 mm 200 mm 230 Holder temperature Normal Increased by 50 °C 290 Holder heat transfer Normal Decreased 315 Temperature below holder Normal Decreased by 350 °C 230 Baking-zone position is defined as the mean vertical distance between the 500 °C isotherm and the lower end of the contact clamps, measured at the fin edge and electrode surface. The standard position is 270 mm.

Thermal Stresses Breakage in the baked paste as the result of thermal stresses is most commonly the reason for electrode damage. Therefore, nearly all the fractures occur in connection with furnace shut-downs of more than 3 to 4 hours. Normally, this kind of breakage occurs horizontally in the electrode and at the furnace-charge level. The present picture is based on collected data from a lot of furnaces. Data from furnace stops with no breakages were also studied for comparison. It is not possible to measure the thermal stresses during an actual operation, so in this case practical experience was used for verification of electrode models. Table 2 demonstrates the effect, on calculated tensile stresses after a furnace stop, of the changing of parameters in both stationary and transient conditions. The critical electrode part is in the centre of the electrode and at the charge level. The figures show the importance of having a 'cold* electrode before shut-down—low electrode current density and high slipping rate. But the stresses are also influenced by reasonable changes in the furnace operation, by the process conditions, and by the properties of the electrode materials.

115 TABLE 2

Calculations of thermal stresses in an electrode after Ike stopping of a ferrosilicon furnace

Electrode conditions Standard Changed to Max. tensile stress* Current 130 kA HOkA 2,8 Slipping rate per 24 h 0,6 m 1,0m 3,1 Electrode diameter 1,7m 1,55 m 4,5 Recovery time 6h 3h 3,3 Furnace stop 6h 12h 4,8 Furnace type Open Closed 3,7 Thermal expansion Normal Decreased 3,0 Modulus of elasticity Normal Decreased 2,6 * The standard tensile stress is 3,5 N/mm

116 SESSION G3 Plasma Technology

Paper C3/1

METALLURGICAL REACTION PHILOSOPHIES OF TRANSFERRED-ARC PLASMA FURNACES

by

K.U. Maske* and K.J. Reidf "Council for Mineral Technology, South Africa tUniversity of Minnesota, U.S.A. Essentially, there are two types of transferred-arc plasma furnaces: those using ancillary anodes, and those using metal-bath anodes. Their potential advantages lie in the large number of ways in which energy can be transferred effectively to different stages of a reduction-reaction sequence. Although the underlying principles of arc physics and chromium chemistry apply to both types of reactors, the differences in their plasma configuration call for two distinct philosophies to explain the metallurgical reactions, which are 'in flight' or 'bath'. Chromite-smelting trials in the lOOkW transferred-arc molten-anode plasma furnace at the Council for Mineral Technology (Mintek) revealed distinct smelting zones. These zones, named arc-attachment, slag-metal bath, and frozen-heel, are evaluated briefly in terms of the parameters affecting the optimum smelting of chromite via the bath- reaction sequence dictated by this plasma configuration. The transferred-plasma arc is preferably attached to particles of reducing agent floating on the bath of molten slag-metal. The intense dissipation of energy at the localized arc attachment provides the energy necessary to sustain the status quo of the bath of molten slag and metal, and causes a loss of reducing agent from the system. Any mechanism that increases the energy dissipated at the anodic-arc attachment, or that favours arc attachment to the particles of reducing agent, increases the rate at which this agent is lost from the system. The bulk average temperature of the zone of the slag-metal bath, and thus the reaction temperature, is dictated by the composition of the slag. The intense energy dissipated at the arc-attachment zone provides the energy necessary to sustain a bath of molten slag-metal, even at extremely high liquidus temperatures. However, the size of the bath, and thus the smelting zone, varies, depending on the energy and sensible heat required, and on the heat losses from the bath. The frozen-heel zone is essentially the interface between the slag-metal bath and the surrounding refractories, and provides necessary protection for the anodes embedded in the hearth refractories, and for the refractories themselves. The transferred-arc plasma furnace employing an ancillary anode can deliver reactants direct into the plasma zone at a controlled rate. A preliminary investigation of chromite reduction at the Minerals Resources Research Center (MRRC) in the U.S.A. revealed that reduction levels of over 50 per cent for iron, and 7 per cent for chromium, can

117 ) be achieved within the plasma zone, i.e. 'in flight'. This preliminary work precipitated ! the need for an investigation of chromite smelting in this type of furnace, because both ! ! 'in-flight' and 'bath' reactions can be achieved in this plasma configuration. This work i is to be undertaken at MRRC. Special emphasis will be placed on the particular j utilization of the plasma arc, its effect on the chromite-reduction mechanism, and the • metallurgical approach needed to realize optimum conditions for the smelting of ferrochromium. Paper C3/2

THE DISSIPATION OF ENERGY IN d.c. TRANSFERRED-ARC PLASMA SYSTEMS AND THE CONSEQUENCES FOR THE PRODUCTION OF FERRO-ALLOYS

by

t A.B. Stewart ! Council for Mineral Technology South Africa

The production of ferro-alloys in the World today is largely carried out in the submerged- arc furnace, which has become a high-powered, reliable production tool. The technology of submerged-arc furnaces is well established and has attained a high degree of sophistication. However, these furnaces exhibit distinct disadvantages, particularly with . regard to the treatment of fine raw material, and there has been great interest recently in the possible application of plasma-arc furnaces to the production of ferro-alloys. : A major contender in this application is the d.c. transferred-arc plasma system. This system employs a conventional open-bath furnace with a direct-current arc set up between an electrode and the bath of molten material in the furnace. (Multiple-electrode systems are also used.) The electrode is always of negative polarity (the cathode), and the bath is of positive polarity (the anode). The electrode is either of the conventional graphite type, which is consumed during operation, or of the non-consumable, water- cooled, tungsten-cathode type. An important factor in the operation of the d.c. transferred-arc furnace is the mode of energy dissipation within the furnace. Numerous experiments have been reported in the literature regarding the dissipation of energy at the electrode of an arc. However, no work has been reported on the overall dissipation of energy, particularly for long arcs (greater than 10 cm). A calorimeter system consisting of a cylindrical furnace made from water-cooled copper panels was constructed for experiments on the dissipation of energy within a furnace. The roof and base of the furnace were water-cooled, and a separate water-cooled anode was installed in the centre of the base. A d.c. transferred arc was set up between the anode and an electrode that was introduced vertically through the roof of the furnace, both graphite and water-cooled electrodes being employed. All the water circuits were monitored separately for flow and temperature difference, and the input of electrical energy was measured so that an overall energy balance could

118 105

be maintained. Experiments were carried out on a power supply of 100 kVA with arc lengths of up to 20 cm, currents up to 700 A, and voltages up to 120 V. These experiments established the overall mode of energy dissipation. Further experiments were conducted on the use of a larger power supply to establish longer arcs and operate at higher voltages and currents. In these, only the input of energy and its dissipation at the anode were monitored. The following are the important conclusions from the experimental work. (1) There is a large dissipation of energy at the anode (up to 80 per cent). This is largely because the energy developed in the arc column is directed downwards onto the anode. (2) The major factors affecting the relative proportion of energy dissipation at the anode are as follows. • A water-cooled tungsten torch causes a higher energy than the carbon-electrode type. • A higher arc current increases this proportion of energy at the anode. • A longer arc decreases the proportion of energy at the anode. • Nitrogen, being a diatomic gas, causes'a higher transfer of energy to the anode than does argon, a monatomic gas. (3) The dissipation of energy from the arc column (radiation energy) is relatively small. As long as the arc column is well collimated and directed downwards onto the anode (or bath of the furnace), the exposed arc column does not result in v high wear of the refractories. The major problems with refractories in a transferred-arc furnace are slag attack on the walls of the furnace, and radiation to the roof of the furnace from the hot anode region and the molten bath.

119 106

SESSION C4 i Plasma Processes * \ : Paper C4/1

| NON-FERROUS METALS EXTRACTION USING TETRONICS PLASMA ! SYSTEM, WITH PARTICULAR EMPHASIS ON ZINC AND LEAD

' by

D.G. Page and C.P. Heanley ... .. j Tetronics Research and Development Co. Ltd England The history of the production of non-ferrous metals shows that a variety of possible techniques can arrive at a similar product. Economies of time or place, feedstock variations, or product constraints have resulted in one method or another being preferred j by different operating companies. For a development programme to succeed in producing L~ an improved process, a versatile approach is needed to encompass the diverse and complex nature of the industry's requirements. Both raw-material supply and the market . for finished products are often unique to a particular producer, and therefore the studies '• and development work must either be applied to the solving of problems common to i most cases, or be tailored to suit one particular situation. ; The Tetronics Plasma System, based on a direct-current transferred arc using a 1 , tungsten cathode that is not consumed during processing, is well tried and tested up to a power of several megawatts. At their pilot facility in Faringdon, Oxfordshire, in the U.K., Tetronics Research and Development Co. Ltd have studied the extraction of various non-ferrous metals from both primary and secondary sources of raw materials. I Testwork has been undertaken at powers of approximately 100 to 200 kW, aimed, for example, at the removal and collection of tin and silver from complex ores and secondary * - residues, to pilot-plant operation in excess of 1 MW, treating dusts arising from steel production. This wide experience, together with the opportunity to operate the experimental facility at Faringdon at powers up to 2 MW continuously, puts Tetronics in a unique position to assess the possible benefits of applying plasma to the extraction of non-ferrous metals, particularly zinc and lead. The following issues are identified to support the proposition that the development of new processes based on the Tetronics Plasma System is a viable commercial concept. (1) Many existing production facilities are outdated, and either pressure from environmental protectionism or an increased market demand will provide an opportunity, if not a necessity, for industry to consider new processes. (2) The market for metals such as lead and zinc is fairly stable in the long term, and plans for a development programme with the recovery of capital expenditure within a reasonable length of time can therefore be made. (3) Many non-ferrous metals are available in commercial quantities from secondary wastes and residues, and a versatile, capitally inexpensive process capable of recovering these metals could be the basis of a profitable exercise.

120 Several aspects of the Tetronics Plasma System have been identified that offer significant benefits in extracting lead and zinc from both oxide and sulphide states; these are summarized as follows. (a) The fine feeds arising in the industry can be fed directly to the furnace without agglomeration, pelletizing, or sintering. (b) The oxidizing or reducing state of the furnace atmosphere can be closely controlled because no carbon electrodes are employed and the furnace is sealed. (c) The gas throughput in the system is small in comparison with other techniques, thereby reducing problems oi' gas cleaning and material carry-over. (d) Any required temperature in the furnace can be closely maintained. (e) Elemental sulphur, rather than sulphur dioxide, can be produced from sulphide feedstocks. •I (f) The process shows good fuming characteristics, the slag depth and chemistry being I chosen and controlled to suit the process rather than the equipment. (g) The electrodes are not consumed during processing. (h) Exacting product specifications can be achieved by utilizing the ability to control the process chemistry independent of the heat source. (i) No barriers exist in the scaling-up of existing process equipment to the level ~~ required for commercial operation. The more dynamic and incisive regime of the plasma arc often leads to improved thermal effects or reactions, which need to be properly investigated and quantified if the process is to operate to maximum advantage. An analogy of applying the Tetronics Plasma System could be that of handing a very sharp scalpel to a surgeon who has previously been using a blunt knife. More useful and effective work is possible, but some learning is needed to control the equipment. Testwork in support of this widely based development programme to extract non- ferrous metals continues, with the objective of determining the design criteria necessary to construct commercial plants.

Paper C4/2

THE APPLICATION OF TRANSFERRED-ARC PLASMA TO THE MELTING OF METAL FINES

by

L.B. McRae, N.A. Barcza, and T.R. Curr Council for Mineral Technology ; South Africa ' Although the melting of high-value metal fines, such as titanium sponge, has been a feasible process for some time, only recently was attention given to the use of this technology for the melting of materials of lower value. Included in these materials are fines arising from the production of ferro-alloys when the cast slabs of alloys are crushed to the size required by the customer, and directly reduced iron (DRI).

121 ^ The physical characteristics of metal fines usually preclude the use of conventional \ techniques for their remelting, and some other means must be used. Blast furnaces are ; : being phased out in several countries, and steel is being produced to an ever-increasing ' extent from scrap and DRI. This technology is known, but in some instances the ! processing of DRI on its own is becoming desirable. Several factors make the I conventional open-arc furnace less than ideal for this purpose. The use of thermal-plasma for the melting of DRI is of great interest to a country like South Africa, where an additional capacity of 600 kt per annum is being installed. The Council for Mineral Technology (Mintek) has undertaken a research-and- ; development programme to find a suitable process for the melting of ferro-alloy and i other fines. As transferred-arc plasma was considered the most suitable approach for this purpose, a 100 kVA furnace was built, which gave continuous steady-state operation over considerable periods of time. The melting of ferro-alloy fines was evaluated on ; that scale, and the results led to the design of industrial-scale facilities. : Another area in which the use of a transferred-arc plasma proved successful on a small scale is the remelting of silicon fines and scrap. Fines are generated when the cast metal is crushed, and the scrap consists of material that has been recovered from launders, usually being heavily contaminated with slag. Extended tests showed that a high recovery of silicon can be expected from the remelting of scrap with a transferred- arc plasma, and that a high-grade slag-free product can be obtained. Water-cooled plasma torches are at a sufficiently advanced stage of development for them to be eminently suited to the demands of remelting operations in the ferro-alloy industry. A typical large ferrochromium plant can produce some 200 kt of alloy per year, and the amount of fines generated from this would be about 20 kt. Based on an operation of 300 days per year, and a power consumption of 700kW*h/t, a 2MW furnace would be required for the remelting operation. At a predicted voltage of 500 V, a current of only 4000 A would be required, which is well within the capability of a single water-cooled transferred-arc plasma device. The largest producer of ferromanganese in South Africa has a capacity of over 0,5 Mt per year, and up to 50 kt of metal fines can be generated. A melting facility to cater for this quantity of material would require a rating of about 5 MW, based on a reasoning similar to that used above for ferrochromium. If this furnace were operated at 500 V, the amperage required would still be within the capabilities of present-day water-cooled transferred-arc plasma torches, i.e. 10kA. The use of a water-cooled transferred-arc plasma system for the melting of ferro- alloy fines has a number of advantages., the most important of which are the high melting rate obtained because of the good heat transfer when metal fines are fed to a bath of molten metal, and the excellent control of the bath temperature. However, problem areas exist, including the anode connection, which must be carefully maintained, and the roof refractories, which, because of the open bath, experience a high thermal load. Solutions to some of these problems have been proposed and implemented. The melting of DRI is more complex. Because of the high throughput required, single torches are inadequate. Multiple torches can give problems with regard to the interaction of the arcs, although there is a furnace design that overcomes this problem. Another

122 109

method by which this problem can be overcome is the use of alternating current. This technology is at present limited to an amperage of 3,3 kA. Graphite electrodes operating on direct current are an alternative to water-cooled torches, and a 40 MW direct-current arc furnace would be feasible. The advantages of this approach are that the graphite electrode is capable of carrying higher currents than a water-cooled torch and scale-up is therefore less of a problem, and that the use of direct current gives lower electrode wear and greater arc stability. However, on the negative side are the costs involved in the replacement of the electrode as it is consumed, and the possible lower current-carrying capacity of the electrode for direct current than for alternating current.

Paper C4/3

INDUSTRIAL PROCESSES BASED UPON PLASMA TECHNOLOGY

by

S. Santen and J. Skogberg SKF Steel Engineering AB » Sweden I A thermal plasma is characterized by a high temperature level and a high energy density. This makes the plasma extremely suitable for high-temperature processes. These features are used within SKF Steel in the development of industrial processes based upon plasma technology. The heart of all the SKF plasma processes is the plasma generator PLASMATECH. This system was developed for an industrial environment covering power levels from 1 to 10 MW. The thermal efficiency of a plasma generator is Mgh, normally 85 to 90 per cent. In addition, very high gas enthalpies can be readied, typical values being 4 to 8kW*h/m3. The ability of a plasma to accumulate large amounts of heat energy within a small volume is important. This energy can easily be transferred and utilized in a subsequent high-temperature process. The process for the direct reduction of iron ore to sponge iron is called PLASMARED. The iron ore is reduced by a mixture of carbon monoxide and hyd:ogen in a continuously operating shaft furnace. This process has been in industrial operation since 1981 in I SKF's PLASMARED plant in Sweden. The reduction gas for the PLASMARED process is supplied from a plasma-based gasifier, in which the carbon monoxide- hydrogen mixture is generated from a fuel, such as coal, heavy fuel oil, natural gas, charcoal, or peat. The coal is gasified with steam and oxygen, and a minor part of the heat energy needed is supplied by hot gas from plasma generators. This gives an extra degree of freedom in the gasification process. It becomes possible to keep the contents of carbon dioxide and water in the product gas at very low levels. In addition, the gasification temperature can be held sufficiently high to make the ash in the coal melt ;; and be removed in the form of a liquid slag. This technology can be used for the regular

123 production of fuel gas, reduction gas, and synthesis gas. A plasma-based coal gasifier ) has been in industrial use since May 1982 at SKF's PLASMARED plant. > PLASMASMELT is a two-stage smelting reduction process with prereduction in a ; fluidized bed, and smelting reduction of the prereduced material in a shaft furnace. j The heart of the process is the smelting zone, which consists of a cavity created in a | coke column in front of the plasma generator. The plasma generators are located in • the lower part of the coke-filled shaft furnace. In the smelting zone, the prereduced iron ore fines are reduced instantly, coal being mainly used as reductant. The necessary heat energy for the reactions is supplied by the plasma generators. The reaction products are hot reducing gas, liquid metal, and slag. The hot metal and the slag are collected in the bottom part of the shaft. The hot reducing gas is withdrawn from the top of the furnace and is used for prereduction of the iron ore fines in a system of fluidized beds. The process provides for optimal energy utilization, with no excess energy leaving the process. The PLASMASMELT concept has been used for the development of a family of processes for the production of ferrochromium, zinc, and other metals and alloys. The PLASMACHROME process is a one-stage process in which fine chromite materials are injected together with coal into a coke-filled shaft furnace. (The i PLASMACHROME technology is based upon the PLASMASMELT process.) The PLASMACHROME process was developed for production of high-carbon ferrochromium. A wide range of raw materials such as concentrates, friable ores, and run-of-mine fines are used without agglomeration. The flexibility of the PLASMACHROME process in selecting raw materials is gaining in importance, because rich deposits of good lumpy chromite ores are becoming scarce. Lower-grade materials must be used to meet the demand for chromium. The PLASMACHROME process has shown its viability in the processing of these fine materials. The PLASM ADUST process is used for the recovery of metals — such as zinc, lead, iron or iron containing chromium, nickel, and molybdenum — from waste oxides in baghouse dust. Such dust is frequently deposited in waste dumps, where heavy metals from the dust can pollute the ground water and cause serious environmental problems. The first commercial PLASM ADUST plant will be commissioned in 1984 by Scan- Dust AB at Landskrona, Sweden. This plant will be the first of its kind in the World. Intensive development work during the seventies resulted in several commercial processes based upon plasma technology. These new industrial processes are energy- efficient and highly productive, and can be applied in small-scale operation.

124 Ill

Paper C4/4 SMELTING REDUCTION OF COMPOSITE CHROMITE PELLETS BY PLASMA

by

S. Kouroki, K. Morita, and N. Sano University of Tokyo Japan The investigation described was undertaken to find the most reasonable route for the production of stainless steel from chromium and iron minerals other than by the conventional process. The common guiding principle was that, if the chromium content in the product was to be lowered to around 20 per cent, the reduction could proceed more rapidly and efficiently because of the lower chromium activity. Reduction of Chromite and Hematite by Carbon A briquette of 1,0 g was made by the pressing of a mixture of chromite, hematite, and graphite (or natural minerals and coke powders), and was heated continuously at a rate of 400 °C per hour under a neutral atmosphere. The change in mass on reduction * was recorded by a thermobalance. In some cases, the reduction was interrupted and I the sample was submitted to optical microscopy, and to X-ray-diffraction, electron- * microprobe, and chemical analysis. At a chromium-to-iron ratio (by mass) of 0,25, the iron reduction was completed at 1170°C to form iron-carbon melts before the chromium reduction began, which ended at 1380 °C. Figure 1 shows the reduction behaviour of chromium for various combinations of iron and chromium. In the presence of excess iron, the chromium reduction occurred at lower temperatures (50 to 100 °C lower) than without it, either with synthetic samples or with natural minerals. The formation of a molten iron- chromium-carbon phase is presumed to enhance the reduction of chromium, but the presence of gangue in natural minerals seems to hinder the diffusion of chromium in the solid oxide and its dissolution in the melt.

Smelting Reduction of Chromite by Plasma A transferred-arc direct-current plasma furnace with a power input of 15 kW was used for the smelting of 2,0 kg of composite pellets of chromite with or without coke and ; iron ore. The pellets were 5 to 8 mm in diameter, and were smelted in a magnesia crucible 10 cm in diameter. Induction stirring was used to improve the reduction. Firstly, three different charges were compared: (1) composite pellets of chromium ore + iron ore + coke, (2) composite pellets of chromium ore + coke to make a metal bath high in chromium, followed by the charging of composite pellets of iron ore + coke, and (3) the reverse order of (2). In all cases, the amount of each constituent remained unchanged where the mass of coke fed was 11 per cent more than was needed to reduce the iron oxide and chromite with carbon monoxide as a product. The chromium recovery from charge (3) was found

125 1.0

Heating rate 400°C/h O O Synthesized FeO

0,8 • • Synthesized FeOCr2Oj hematite mixture (Cr/Fe 0,25) .—O Chromite ore Chromite ore- § iron ore mixture (Cr/Fe = 0 0,6 I•s U I* 73 e o 0,4

0.2

0 1000 1200 1400 Temperature (°C)

FIGURE 1. Fractional reduction of chromium with various composite pellets

to be most favourable, although 13,3 per cent of the chromium remained in the final slag as solid MgO'Cr2O3. In sharp contrast to charges (1) and (2), 40 per cent of the phosphorus input was found to have been removed from the system, probably because phosphorus vapour from the iron ore at the first stage was carried away by the large amount of carbon monoxide that evolved violently when the iron ore was reduced. As a result of these findings, all the subsequent experiments were carried out in such a way that the composite pellets of chromite ore and coke were charged to a bath of molten iron. In that case, the main part of the chromium reduction seemed to occur in the molten slag with suspended coke, rather than with carbon in the metal bath,, and the minimum carbon level in the bath needed for the coke to be suspended was found to be around 4 per cent. The composition of the slag was another important factor since the rate-determining step is likely to be the dissolution of solid MgO'Cr2O3 into molten slags. The optimal composition was around 10 per cent CaO, 40 per cent SiO2, 30 per cent MgO, 15 per cent AI2O3, which was made by the addition of SiO2 and CaO to the original minerals. An increase in MgO was unavoidable owing to its dissolution from a crucible.

126 Figure 2 shows the reduction behaviour when the metal bath is first saturated with carbon to make an 18 per cent and then a 36 per cent chromium-containing alloy. Silicon and sulphur precisely followed the reducing condition resulting from the addition of excess coke. The composition of the metal after reduction was typically 18 per cent chromium, 4,65 per cent carbon, 1,3 per cent silicon, 0,004 per cent sulphur, 0,037 per cent phosphorus with a chromium yield of 97,5 per cent. Significant gaseous desulphurization was a feature of the smelting reduction. Similar results were obtained when coke was replaced by coal, except that more gas and dust formed. It is concluded that the making of molten iron-chromium-carbon alloys directly from minerals is technically feasible, although the alloys must subsequently be decarburized before they can be used in the production of stainless steel.

Additional charge of cokes

40 60 80 100 120 140 160 Time (minute)

FIGURE 2. Changes in Cr, Si, C, P, and S concentrations in metal with time on smelting reduction of chromite by plasma

127 SESSION C5 Plasma Facilities

Paper C5/1

THE 3,2 MVA PLASMA FACILITY AT MINTEK

by

T.R. Curr, K.C. Nicol, J.F. Mooney, A.B. Stewart, and N.A. Barcza Council for Mineral Technology South Africa The Council for Mineral Technology (Mintek) has been interested in the application of thermal plasma to pyrometallurgy for some time and, in particular, to the production of ferro-alloys. In South Africa, there is a large installed capacity of submerged-arc electric furnaces (for ferro-alloys), as well as conventional open-arc steelmaking furnaces, to which the technology of thermal plasma is relevant. Mintek was especially aware of the potential advantages of thermal plasmas for the direct treatment of chromite-ore fines, and therefore enthusiastically undertook to investigate the plasma smelting of ferrochromium in the United Kingdom when approached to do so by a large local producer. The success of these trials, which used a transferred-arc system, encouraged Mintek to begin setting up its own pilot plant to support both the initial effort in the ferrochromium industry and the further application of thermal plasma to local materials. It was decided that both a small-scale furnace (for the investigation of the chemistry of various processes) and a larger-scale facility (for scale-up purposes) would be installed. The small-scale (100 kVA) facility has been discussed elsewhere, and this paper concentrates on the large (3,2 MVA) facility and gives an evaluation of its performance. This capacity was decided on as providing a thirty-fold scale-up factor on the lOOkVA furnace, as well as being sufficient to reach a thermal efficiency that is comparable with that used in industrial practice. A direct-current transferred-arc configuration was selected, partly because of the success achieved during the overseas testwork with this configuration, and partly because a number of potential advantages had become apparent from tests conducted on the lOOkVA furnace. The high degree of metallurgical control possible with the open-bath transferred-arc system, in which the feed is reacted to completion in a thin layer on the surface of the liquid bath, allows a vast increase in the number of processes that are at least potentially possible in a plasma furnace over those in submerged-arc furnace practice. Secondly, the power-supply requirements for the transferred-arc approach are far more compatible with the power supplies for existing submerged-arc furnaces than with the non-transferred arc system. Once the capacity of the system had been set, the furnace and power supply together with the associated support and service systems were designed to allow for maximum flexibility. The whole system can be divided into five distinct sections: furnace, power supply, feed system, services, and product handling. The services include cooling water,

128 11O

plasma and purge gases, hydraulics, and steam. The product handling includes both the hot-metal handling and the off-gas system. The physical dimensions of the furnace were designed on the experience gained from trials locally (at a level of lOOkVA) and overseas (at a level of 1400 kVA). Several innovative features were incorporated in the design. These include die use of two horizontal furnace-gas off-take ducts, a hydraulically driven damper system in these ducts to control the furnace pressure, and vertical stainless-steel anodes situated in the hearth of the furnace. The power supply was designed to allow for maximum flexibility so as to accommodate both transferred and non-transferred arc systems. Two six-pulse thyristor-based units were installed, and these can be run individually, in series, or in parallel to cater for the voltage and current needs of the various arc systems. Since the main advantage of the transferred-arc type of system is its controllability, considerable design effort was used to produce a feed system that would not only feed fine material but also allow the feedrate into the furnace to be controlled. Although no innovative hot-product handling methods are employed, a unique off-gas gas-cleaning system employing combustion chambers, watercooled quench chambers, and a hydrosonic-venturi scrubber unit has been implemented. Since the entire unit was designed for maximum flexibility, a programmable logic controller was used to perform both the interlocking and the control functions. The transferred-plasma arc system allows wide control of the variables such as arc power input, feedrate, and furnace pressure, as well as the direct measurement of bath temperature. The paper concludes with an evaluation of the facility as a whole, with special emphasis on its innovative features.

Paper C5/2 AN EXTENDED ARC REACTOR, USING GRAPHITE ELECTRODES

by

C.B. Alcock, A. McLean, and I.D. Sommerville University of Toronto Canada The introduction of argon into the discharge volume of a carbon arc has long been known to lead to a discharge that is more stable than that struck in air. If, however, the argon is introduced through small-diameter holes pierced axiaily through the two graphite electrodes of a single-phase system, it has been found that a stable and extendible arc is formed. This discharge is quieter, both electrically and acoustically, than the normal arc, and provides a more uniform and diffuse energy source. The quantity of argon that is used to stabilize the extended arc is extremely modest, and thus the system presents attractive economic aspects for industrial applications. The extended arc has been coupled with a rotary preheater, a freefall 'flight tube', and a conventional furnace configuration to constitute a reactor in which a number of oxide reduction studies have been carried out. A wide range of steel-plant waste oxides,

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and fines producing ferrochromium, ferromanganese, and ferrovanadium molten alloys, have been treated in a three-phase extended arc reactor on a power scale up to 100 kVA. The flexibility of the reactor, which makes possible the reduction of mixtures of oxides having a wide dispersion in average particle sizes, and by diverse types of reductants, from charcoal to anthracite, is enhanced by variation of the argon flowrate through each electrode. This latter makes it possible to vary the energy input at each pair of electrodes independently. The countercurrent flow of gases and solids optimizes heat exchange, and gases leave the experimental furnace at temperatures less than 100 °C. Feeding rates of oxide, slag-making material, and reductant of 1 kg per minute are easily achieved in the reactor. In a number of studies, some of which are referred to above, it has been demonstrated that in-flight reduction occurs only of those oxides having a relatively low stability. For example, FeO, NiO, and the more stable oxides must be reduced by the conventional slag-metal reaction, where the kinetic reducing capacity of the system is much higher than that of the gas phase. Finally, because the arc is extended during reduction, the carbon electrodes play a negligible chemical role in the reduction reaction, and it is possible to produce low- carbon liquid ferrous alloys by direct reduction of a suitable mixture of oxides. These alloys can be made with very low sulphur and phosphorus contents. This is because the slag composition can be chosen to optimize the elimination of these elements without the parallel consideration of the electrical properties of the slag, which is a very important limiting factor in submerged-arc operation using graphite electrodes. Alternatively, if a graphite crucible is used and liquid iron saturated with carbon is a satisfactory product, zinc and lead can be evaporated from the liquid metal and slag system during the treatment of arc-furnace baghouse dusts.

Paper C5/3

THE DEVELOPMENT OF A HIGH-POWER ARC HEATER—A PROGRESS REPORT

by

A.L. Hare Tioxide UK Ltd England Derivatives of the Schoenherr arc heater have become the process industries' standard design for plasma gas heating. This type of arc heater can be used either as a plasma reactor or as an efficient ultra-high-temperature gas heater. In plasma-reactor applications, its use is normally restricted to homogeneous gas-phase synthesis such as nitrogen fixation or hydrocarbon cracking. When used as a gas heater, the high plasma-jet momentum makes it suitable for a wide variety of rapid heating and mixing processes, including heterogeneous reactions such as the manufacture of pigmentary titanium dioxide.

130 117

The reasons for the popularity of the Schoenherr design are that it can be operated on either alternating- or direct-current power; it is highly efficient with more than 80 per cent of the applied electrical energy converted into heat. Furthermore, the basic configuration allows the use of simple geometric electrode shapes and a robust mechanical design. At present, units up to a power rating of 10 MW are available for heating most gases. However, for the process industries the basic design has certain inherent limitations. To obtain powers greater than 10 MW and high specific gas enthalpies, it is necessary to use arc currents significantly greater than 1000 A. It has been demonstrated that, at arc currents over 1000 A, the rates of electrode erosion increase very much more rapidly than the increased power output obtained. As a result, plant downtime increases disproportionately and, in certain specialist applications, product contamination by i" electrode material reaches unacceptable concentrations. i Several arc-heater designs have been reported which could overcome these limitations, I but for a variety of reasons these designs do not appear suitable for the process industries. An arc-heater design which is compatible with process-industry requirements has been j developed. The key feature of this design is the use of multiple switchable inter-electrode sections. The development of a reliable switching system proved to be more difficult .TM_ than expected. The provision of a controlled gas flowrate between sections was essential for stable operation. The number and size of the inter-electrode sections required depends I ' on the particular application. Various factors have to be taken into account, including 1 : the specific enthalpy required and the physical space available. ! Once (these problems had been overcome, the objectives of the initial programme • were achieved. Evaluation of the full potential of this arc-heater design was not possible l- because of the limited voltage rating of the power supply. By comparison with 1 conventional arc-heater performance, it is believed that 20 MW plasma-gas heaters, suitable for process use, are now practicable since electrode lives of several hundred . < hours should be possible with most process gases. t Paper C5/4 THREE-PHASE a.c. PLASMA MELTING: STATE OF DEVELOPMENT AND POTENTIAL APPLICATIONS

by t

H.J. Bebber, J. Hartwig, D. Neuschutz, and H.-O. Rossner

Fried. Krupp GmbH, Krupp Forschungsinstitut ; West Germany For several years, Krupp Research Institute has been developing three-phase alternating- current plasma-melting furnaces. At present, a 31 pilot furnace is being operated with alternating-current torches at up to 3,3 kA. Plasma torches for 6kA are under development. A 101 alternating-current plasma furnace rated at 10 MW electric power is being built at Krupp Stahl AG for scrap melting, and will be commissioned at the end of 1984.

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Several advantages are attributed to direct-current plasma furnaces of the Freital type ) as compared with conventional electric-arc furnaces: ! - the use of argon as the plasma gas and the easy sealing of plasma furnaces lead I to a relatively inert furnace atmosphere and, consequently, to a high yield of alloying • elements and iron; 1 - the stability of the plasma arc results in low-noise operation and drastically reduced '• flicker; - the carburization of the melt by electrode graphite is avoided; - the operating costs are decreased, since the costs of argon are lower than those of graphite. The incentive to develop a three-phase alternating-current plasma furnace at Krupp j was based on the assumptions that it should be possible to design the torches and power- supply system to obtain stable plasma arcs that would combine the advantages of both direct-current plasma arcs and alternating-current electric-arc furnaces. It seemed particularly desirable to introduce the torches through the furnace roof as in conventional electric-arc furnaces, since this positioning makes it possible to charge low-density scrap and to melt down the scrap from inside out, thus minimizing thermal- ! lining wear. With direct-current plasma, for all cases where more than one torch is L. needed, these torches are introduced into the furnace laterally since in vertical positioning the arcs converge above the bath, leading to inefficient heat transfer and distribution. Moreover, the bottom electrode is avoided in three-phase alternating-current operation. For these reasons, it seemed worthwhile to focus further developmental efforts on the three-phase alternating-current plasma furnaces. The research programme at the Krupp Research Institute had to include the development, not only of the furnace and suitable metallurgical operating modes, but also of the plasma torches and the electrical power-supply system, since none of these components is commercially available or can readily be transferred from existing direct- current plasma experience. It should be pointed out that plasma torches suitable for application in steelmaking have transferred arcs with one electrode in the torch, and the other electrode in the furnace bottom (direct current) or in the counter torch (three- phase alternating current). Plasma torches with non-transferred arcs, which are commonly used for gas heating, welding, or spraying, have a considerably lower thermal efficiency. The 31 three-phase alternating-current pilot plasma furnace started operating at the Krupp Research Institute in 1982. The furnace has an outer diameter of 2,3 m, a magnesite lining of the wall and hearth, a water-cooled, alumina-lined roof, and a platform carrying the torch supporting arms, which permit the torches to be moved up and down and to be inclined by 20°. The plasma torches used in this furnace have been run at 3,3 kA. With arc lengths of up to 700 mm, as achieved during the melt-down periods, a power rating of 1,3 MW can be reached per torch. The Krupp plasma torch has a hot, tungsten-based electrode. Argon flowing from the annular gap between the electrode and the surrounding nozzle keeps the electrode from being oxidized and serves to stabilize the plasma arc. Argon consumption is 8mVh per torch. An auxiliary arc is used only for a very short period to ignite the main arc, and is turned off after ignition.

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The power-supply system for the 3t plasma furnace is rated at 6kA and 575 V maximum, but for external reasons the maximum effective power is limited to 1,5 MW. The test melts carried out so far in this pilot furnace have demonstrated the usefulness of the power supply, the torches, and the furnace design. The melting of scrap has been investigated in detail, including questions of arc stability, noise emission, and thermal efficiency. One aim of the plasma-furnace development work at Krupp is a melting unit not only for high-alloy steel but also for tonnage steel. Efforts must be made to keep the costs of electrodes, nozzles, and plasma gas lower than the graphite costs for electric- arc furnaces. Once this is achieved, die advantages of reduced noise emission, less flicker, and a higher metal yield for roughly the same initial cost should provide incentive enough to operate alternating-current plasma furnaces. /'- Another interesting field of application of plasma technology is in ladle furnaces. , Compared with graphite electrodes, plasma torches offer the advantages of a more inert i atmosphere by the plasma-gas argon (no danger of nitrogen pick-up), of no carburization, ! of a lighter construction of the electrode supporting arms, and possibly of lower lining j wear because the arcs can be positioned closer to the ladle axis. i Furthermore, Krupp's plasma technology holds potential for the melting of metallic J_. fines, the smelting of ore fines, and various other applications in non-ferrous metallurgy. I

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SESSION C6 Mineral and Material Science in Pyrometallurgy i ] Paper C6/1 i S SLAG-METAL EQUILIBRIA INVOLVING CHROMIUM AS A COMPONENT

1

A. Muan The Pennsylvania State University U.S.A.

The paper deals with high-temperature equilibrium relations in oxide and silicate systems involving chromium oxide as a component. Under strongly reducing conditions, such as exist for instance when silicate melts (slags) are in equilibrium with alloy phases (e.g. Fe-Cr), significant parts of the chromium in the silicate phase may be present as Cr2 +. Such relations are illustrated by phase-equilibrium diagrams which were determined i under controlled, strongly reducing conditions for some key chromium oxide-containing systems, and by thermodynamic data for phases in the systems Fe-Cr-O and CaO- , chromium oxide-SiO2. Examples of the former are shown by the systems MgO- chromium oxide-SiO2 and CaO-chromium oxide-SiO2 as determined under different oxygen potentials, for instance air and in equilibrium with metallic chromium. It has been shown that drastic decreases in liquidus temperatures and increases in the solubility of chromium oxide in the silicate liquid result under the strongly reducing conditions imposed by the presence of metallic chromium as compared with the situation existing in air. Activity-composition relations in Fe3O4-'Cr3O4' solid solutions have been determined from equilibrium measurements involving this phase, Fe-Cr alloys, and a gas phase of known oxygen potentials obtained by mixing CO2 and H2 gases in controlled proportions. The thermodynamic behaviour of the solution between the FeCr2O4 composition and the hypothetical Cr3O4 end-member can be well described by the substitution of Cr2 * for Fez + in tetrahedral sites of the iron chromite with increasing chromium-to-iron ratio of the solid solution. Activity-composition relations of chromium oxide in CaO-chromium oxide-SiO2 melts have been determined by equilibrating such melts with Mo-Cr alloys of known activity-composition relations in atmospheres of known oxygen potential obtained by mixing gases (CO2 and H2 or H2O and H2) in controlled proportions. Activities of Cr2O3 can be calculated from the chromium activities of the alloy and known oxygen i fugacities of the gas phase by the relationship

. I ^2 V/2 aCrO, = aCr 5 V) '

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where r is the H2/H2O ratio for the coexistence of alloy and liquid silicate, and r0 is the value of H2/H2O for the coexistence of pure Cr2O3 and pure Cr metal at the same temperature. Combining data for the above equilibrium with data from the system Cr-O permits calculation of the activities of CrO in silicate melts. Comparison of the data obtained with those reported previously for activities of FeO in the system CaO-FeO- SiO2 indicate a similarity in the behaviour of CrO with that of FeO in silicate melts. The distribution of chromium, iron, and silicon between alloy and slag phase was determined recently for the system Fe-Cr-Si-O. Mixtures of various compositions of slags in this model system and Fe-Cr-Si alloys of chromium contents up to 30 per cent (by mass) and silicon contents up to 2 per cent (by mass) were equilibrated under an argon atmosphere in the temperature range of 1600 and 1700 °C. Fused silica was used i to contain the melts in this study. Thermodynamic relations derived from the :' experimental data on the equilibrium between slag and alloy indicate that chromium 1 in the slag is predominantly present in the divalent state. The solubility of silica in slags ; that are rich in chromium oxide and saturated in silica is approximately constant at , constant temperature. The effects of chromium and temperature on the solubility of I silicon in the alloy have been determined. The interaction parameter of Cr on Si («$;) { in liquid iron at 1600 °C is close to zero. Additional studies of the distribution of chromium between slag and alloy phases were made recently for slags in the system CaO-MgO-chromium oxide-SiO2. These equilibrations were carried out in MgO crucibles. The data show that low FeO contents of the slag, high Si and Cr contents of the alloys, and low temperature favour the recovery of chromium from slag into alloy. Increasing basicity reduces the loss of chromium to slag effectively when the basicity is relatively low, but the beneficial effect levels off when the basicity reaches a value of approximately 1,8.

Paper C6/2 THE MINERALOGY OF THE REDUCTION OF CHROMIUM ORE IN A PLASMA FURNACE by A. Wedepohl Council for Mineral Technology South Africa Two possible methods for the processing of ores that occur as fines are (1) agglomeration with subsequent feeding of the pellets or briquettes to a furnace, and (2) direct feeding of the fines to a suitable furnace. For the latter method, a plasma furnace suggests itself as a possible candidate since fines can be fed direct through the hot plasma. Work on an experimental transferred-arc plasma furnace is currently in progress at the Council for Mineral Technology (Mintek) so that the feasibility of the reduction of South African chromite ore by this method can be assessed. The work includes

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optimization of the reduction process by changing of the operating parameters such as geometry, particle size, slag composition, and atmosphere. The composition of the slag can be varied more widely in a plasma furnace than in the traditional submerged- arc furnace, since mo restraints are imposed by required values for viscosity, liquidus temperature,, ? J conductivity. In addition, it is easier for the required low fugacity of oxygen to be adhered to in a plasma, furnace. This is achieved by the maintenance of a slightly positive pressure in rtie furnace, thus preventing any inflow of oxygen from the atmosphere. The fugacity of the oxygen is then controlled by the ratio of carbon dioxide to carbon monoxide within the furnace itself. This, in turn, is kept low by the presence of solid carbon in the system, which promotes the Boudouard reaction in favour of the formation of carbon monoxide:

C + CO2 -» 2CO.

Previous work had shown that, initially, the reduction of chromite can take p»ace by a solid-solid reaction, i.e. solid carbon with solid grains of chromite. However, it is primarily ferrous oxide from the chromite that is reduced to iron metal or carbides in this way. Further reduction of the remaining iron and chromium oxides in the chromite takes place by liquid-solid reduction. This implies that, for the production of chromium carbides from the chromic oxide in chromite, two processes must take place: dissolution and reduction. Both these processes can occur only at very low values of oxygen fugacity. The current set of experiments concern the in-bath, rather than the in-flight, reduction of chromite. Samples of slag and metal from various runs in a plasma furnace were subjected to mineralogical analysis so that the reduction mechanism could be elucidated and the effect of changes in various parameters observed. The efficiency of the dissolution process can be assessed by the presence or absence of partially altered chromite particles. If dissolution of the chromite particles is complete, there is still a possibility that the reduction process has not been completed, and the slag consists of blue glass containing many small, red euhedral crystals of spinel. These colours indicate a high chromium oxide content: CrO in the glass, and Cr2O3 in the spinel. Small metallic blebs consisting of a chromium-rich solid solution of a-(Cr.Fe) are always attached to the spinel crystals. This gives rise to the theory that, while the melt is hot, all the chromium dissolves in the glass phase as Cr2+ but that, as the melt cools, some of the Cr2+ is incorporated as Cr3+ in the spinel crystals and as Cr° in the metal, according to the reaction

z+ 3+ 3Cr - 2Cr + Cr°meta]-

No phases containing large amounts of chromium had been found in work on samples from the dig-out of an arc furnace. The reason for this is thought to be that the reduction rate was faster than the dissolution rate, which was therefore the rate-controlling process, and that very little residual chromium remained dissolved in the slag. Residual chromium

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occurred primarily in the undissolved particles of chromite. However, in a plasma furnace, the dissolution is much faster than in a submerged-arc furnace, owing to the high temperatures that can be achieved and the low fugacities of oxygen, so that reduction is the rate-controlling process. Most of the metallic samples examined consisted of (Cr,Fe)7C3 plus a-(Fe,Cr) solid solution. In exceptional cases, the carbide (Cr,Fe)23C6 was present. It was found that an efficiency of more than 90 per cent was attained in the reduction of Winterveld fines by the experimental transferred-arc plasma furnace under optimized conditions.

Paper C6/3

CHROMITE MINERALOGY AND METALLURGICAL BEHAVIOUR

by

T.R.C. Fernandes University of Zimbabwe Zimbabwe The reduction of chromium ore results in different ferro-alloys, and the carbon contents of these alloys depend on the physical and chemical properties of the ore and on the temperature of the reaction. In the production of high-carbon ferrochromium, a refining ore is used to produce the carbide (Cr,Fe)26C3 with a carbon content of less than 6 per cent. Non-refining ores and ores with a low Cr:Fe ratio produce alloys with approximately 8 per cent carbon. This study investigates the properties that characterize a refining ore, and those that determine its low rate of reduction as compared with non-refining ores that are easily reduced. The reducibilities of chromium ores from Zimbabwe are a measure of the relative ease with which the ores can be reduced to metal, and these vary greatly and show a strong negative correlation with the refractory index. The refractory index is the ratio (Cr2O3 + MgO + Al2O3)/(total Fe us FeO + SiO2). The Cr2O3 and SiO2 contents of the ores also vary systematically with reducibility, but iron oxide and magnesium oxide show very little correlation. The Cr:Fe ratio has a reasonably strong correlation with the reducibility of the ores, and some differences are evident between ores from the Great Dyke and podiform ores. The reasons for these correlations and the mechanisms of the reduction reactions are not clearly understood. The spinel unit cell has 8 cations in tetrahedral sites and 16 in octahedral sites. The proportions of the cations in the unit cell were calculated from electron-microprobe analyses, and the divalent and trivalent iron was calculated on the stoichiometry. Iron can occupy either site, and the distribution, estimated from Mossbauer measurements, is likely to have a significant influence on the properties of the spinel and its rate of reduction. Mossbauer studies of spinels indicate anomalous magnetic behaviour in some samples which are believed to be related to the thermal history of the spinel. X-ray-diffraction

137 data confirm the anomalous nature of some of the spinels. t It is proposed that the anomalous properties are the result of microstates within the j , spinels formed during their crystallization, and influenced perhaps by metamorphism ; resulting in the serpentinization of the rock mass. These microstates are believed to j be due to incipient reduction, and are likely therefore to affect the paths and the : mechanisms of the reduction reactions. Magnetic properties of spinels are a sensitive measure of the redox condition of the spinels, and preliminary measurements of magnetic susceptibilities indicate a correlation with reducibility. Different trends are evident for Great Dyke ores and for podiform ores. The analysis of the cell structure of the spinel molecule permits empirical and \ theoretical studies of potential substitutions within the crystal lattice. This analysis could reveal the mechanisms which control the transformation of the spinel to the alloy.

Paper C6/4

FROM METALLURGICAL GRADE SILICON TO SOLAR GRADE SILICON— 1 THE CHEMISTRY OF DIFFERENT APPROACHES

~ by

T.N. Lung Industrial Technology Research Institute Taiwan

The metallurgical-grade silicon (approximately 98,5 per cent silicon) currently produced on an industrial scale by carbothermic reduction of quartz with carbon in a submerged- arc furnace is perfectly suitable for most uses in alloying and in the preparation of silicone polymers. However, the impurity concentration is still some orders too high to be applied as the precursor in producing solar-grade silicon, in which the upper limit of total impurity has been set at 120 p.p.m. for practical reference. In many countries, much research effort was recently devoted to develop the technology of producing low-cost silicon material directly from metallurgical-grade silicon, which is feasible for terrestrial utilization of solar energy. Metallurgical-grade silicon is the most feasible form of starting material for the manufacture of solar silicon when such factors as material availability, reliability, flexibility, and processing economy are taken into consideration. The major impurities in commercial metallurgical-grade silicon are aluminium, iron, and calcium which may reach a few tenths of a per cent; the other minor impurities include boron, phosphorus, ! chromium, copper, magnesium, manganese, nickel, titanium, and vanadium at : concentrations of up to several hundred parts per million. The impurity elements in metallic silicon can be examined as substitutional, interstitial, and precipitated at grain boundaries according to how they are formed. Based on the chemical processes applied to purify the metallurgical-grade silicon, the technical and economical potentially feasible approaches can be classified into four categories.

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I. Chemical upgrading 1. Reactive gas-blowing of molten silicon 2. Acid leaching of solid silicon II. Conversion process 1. Trichlorosilane synthesis and reductive pyrolysis 2. Silicon tetrachloride hydrogenation and pyrolysis 3. Chlorination and metal reduction III. Electrodeposition 1. Molten-salt electrorefining IV. Directional solidification 1. Czochralski pulling 2. Float-zone refining 3. Heat-exchange method. The fundamental chemical aspects of feasible approaches and technical development to produce low-cost solar-grade silicon feed stocks for terrestrial photovoltaic application, together with the basic economic character of future industrialization, are studied and evaluated. I

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SESSION C7 Non-ferrous Pyrometallurgy

Paper C7/1

EXTRACTION OF COBALT FROM NKANA SMELTER CONVERTER SLAGS

by

E. Moyo, J. Kalusa, and S.L. Kher Zambia Consolidated Copper Mines Ltd Zambia Long before the testwork which forms the subject of this paper was conducted, mattes made from cobalt concentrates and in slag-cleaning operations had been tested in Zambia on a laboratory or pilot scale for cobalt recovery through a roast-leach process followed by electrowinning. Cobalt recoveries in both were mostly below 75 per cent for acid leaching. The tests involving mattes made from concentrates were attended by severe sintering problems during roasting. Just before the present work was undertaken, investigations to study the causes underlying sintering revealed that mattes would sinter during roasting as long as iron and cobalt were present in elemental, unsulphidized form. To ensure complete sulphidization, the Nampundwe pyrite concentrate (32 per cent sulphur) used as a sulphidizer had to be increased till the converter slag and sulphidizer ratio became 1:1. The Nkana converter slags used in these tests contained 2 to 3 per cent copper, 2 to 2,5 per cent cobalt, and 45 to 50 per cent iron. The cobalt in the slags is present mainly as a silicate which is amenable to sulphidization in a reducing atmosphere above 1300 °C. An electric resistance furnace, akin to the furnaces universally employed for slag cleaning, would be ideal for such an application in a plant. In this testwork, however, an oil-fired crucible furnace was employed and salamander crucibles were used. The temperatures attained in the tests were 1370 °C as there were no facilities for positive temperature control. The converter slag obtained from the plant was crushed! to 6 mm and blended with pyrite concentrate containing 32 to 35 per cent sulphur. The melts were conducted in batches consisting of 20 kg of slag, 20 kg of pyrite concentrates, 0,8 kg of Maambacoal, and 2 kg of burnt lime. It took about 3 hours to melt a charge. The molten mass was held at 1370 °C for 2 hours after melt-down, and then poured into smaller crucibles and cooled to ambient temperature. Cobaltiferous mattes containing 3 to 5 per cent copper, 1 to 2 per cent cobalt, 55 to 60 per cent iron, and 28 to 30 per cent sulphur were produced. The mattes obtained were pulverized to 60 per cent passing 74 /im, blended with an equal amount of pyrite concentrates, and roasted in a 100 mm diameter laboratory fluosolids roaster at 700 °C using 100 per cent excess air. An air flowrate of 881/min was maintained at 25 °C and a gauge pressure of 103 kN/m2. The calcines analysed 2 to 3 per cent copper, 0,7 to 1 per cent cobalt, and 51 to 53 per cent iron. They were leached under the conditions prevailing at the Nkana and Chambishi Cobalt Plants, which are shown in Table 1.

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TABLE i

Calcine-leaching Conditions

Temp. Time Plant PH °C h Nkana Cobalt Plant 1,8 85 4,0 Chambishi Cobalt Plant 1,4 80 0,5

The leach recoveries obtained are shown in Table 2.

TABLE 2

Calcine-leaching Recoveries

Sample Nkana Cobalt Plant, % Chambishi Cobalt Plant, % no. Cu Co Fe Cu Co Fe I 1 75,6 74,9 1,6 83,0 80,3 1,7 2 84,3 81,2 1,7 86,1 82,3 1.6 3 80,0 82,5 1,3 87,3 80,5 1.1 4 90,3 83,7 1,1 90,7 85,3 1.2 5 90,9 85,7 1,0 89,2 84,1 0,8

The laboratory roaster treats a bone-dry feed, while the plant roasters are slurry fed. Because air is used in the former as an agent of transport and a fluidizer in addition to its being the source of oxygen supply, the excess air values are inevitably higher than in the plant. Owing to this limitation, the laboratory roaster is useful only for ascertaining whether a material is roastable or not and the possible extent of the iron elimination. Cobalt-plant concetiitrates treated in it show leaching recoveries in fair agreement with the values obtained in the plant. For scale-up purposes to determine plant parameters, however, it is recommended that the tests be repeated in a larger, pilot slurry-fed roaster. As the leach liquor at Nkana would join the regular cobalt-plant stream before the purification and electrowinning stages, this testwork was confined to the pyrometallurgical steps.

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i Paper C7/2

{ SPRINKLE SMELTING DEVELOPMENT j I by T.D. Jackson and J.A. Schneider Dravo Engineers Inc. U.S.A.

' First-generation flash smelting, began when people realized the heat potential contained in the iron and sulphur content of chalcopyrite and the possible conservation of energy. Now, second-generation flash smelting, known as oxygen-sprinkle smelting, is utilizing not only the iron and sulphur of chalcopyrite, but also the antiquated reverberatory furnace to make an efficient flash furnace. t With construction completed at a major copper smelter in Arizona, U.S.A., the data j collected show that the MMBTU requirement per ton of new metal-bearing material - (NMBM) has been drastically reduced by utilizing the iron and sulphur content of the concentrate and by increasing the throughput in the furnace. This paper describes some of the problems encountered during the development of the process, and the oxygen-sprinkler burner. The start-up, and the associated problems and solutions that were a part of it, are discussed. Data collected during the start-up operation include the sulphur dioxide content in the gas stream, bath temperatures, and matte-grade control, as well as copper-slag content, dust loading, and steam production. These data are detailed and discussed. The paper also discusses the capital required by this type of retrofit and whether the process reduces costs to the producer.

142 SESSION C8 Applied Measurements

Paper C8/1

FACTORS AFFECTING THE CORROSIVITY OF UNDERGROUND MINE WATERS

by

R.T. White and A. Higginson Council for Mineral Technology South Africa ! The corrosion of water-carrying equipment is a problem of major financial importance I to the South African mining industry. Most of the water in the mines results from i infiltration, and, owing to the large volumes involved, the water is subjected only to i simple treatment procedures such as clarification and neutralization with lime. The j water is then used in processes and machines, and as feed for refrigeration units. Very ~*~ little information is available on the relationship between the corrosion rates of metals (even of mild steel) and the physical and chemical properties of naturally occurring waters. Samples of water that had been collected underground from eight South African gold mines were subjected to a comprehensive chemical analysis, and their corrosivity on mild steel was measured by extrapolation from Tafel plots and by measurement of the polarization resistance. The variation in the corrosivity of a synthetic mine water with systematic variations in its physical and chemical properties was also investigated, mainly i by use of the experimental technique of loss in mass. Analytical results show that a 'typical' mine water is in the neutral pH range, almost fully aerated, and very hard. The major anionic constituents are sulphate, chloride, and nitrate ions (in decreasing order of concentration), and there are two general linear relations between water-quality parameters: a linear increase in conductance with increasing total content of dissolved solids, and a linear increase in Langelier index with increasing pH. The most important factor in the corrosivity of mild steel was found to be the concentration of dissolved oxygen present. De-aeration results in a large decrease in corrosion rates, especially for mine waters in the neutral pH range. The corrosion rate is independent of pH for pH values above 4 in both aerated and de-aerated mine waters. Below pH values of 4 (corresponding to Langelier index values more negative than - 5), the corrosion rates rises exponentially with decreasing pH. The importance of the concentration of dissolved oxygen in determining the corrosivity of mine waters on mild steel is not, as yet, widely recognized in the mining industry. The large decrease in corrosion rates caused by de-aeration points to the possible use of de-aerators in closed-loop mining systems for the purposes of corrosion control. The cathodic reduction of dissolved oxygen is therefore the major process governing the corrosion rate of mild steel in mine waters in the neutral pH range. Tafel plots

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\ for aerated mine waters show that this reduction process is under diffusion control. The | effects on corrosion rates of variations in stirring rate and temperature are also consistent ! with diffusion as the rate-controlling process. \ An assessment was made of the effect on the corrosivity of a synthetic mine water, | of systematic and separate variations in the concentrations of chloride, sulphate, and : nitrate anions. It was found that a variation in the concentrations of any of these anions had no significant effect on the corrosion rates of mild steel. This invalidates the use of corrosivity indices based upon the concentrations of these anions. However, the ammonium ion was found to have a significant effect in increasing the corrosion rate over a wide range of concentrations.

Paper C8/2

CONTINUOUS MONITORING OF INFORMATION FROM THE INSIDE OF A ROTARY MILL

by

L.A. Vermeulen, M.J. Ohlson de Fine, and F. Shakowski Council for Mineral Technology South Africa Relevant information that is obtained continuously from the inside of a rotary mill can make an important contribution towards an understanding of milling dynamics, and could therefore lead to the efficient control of grinding circuits since the information refers to the grinding unit itself. For example, it is generally believed that mills are most efficient when they are operated at or near the maximum value for power versus load, but in the vast majority of mills the load is not measured continuously and the following are not known: the mean density of the load, the relationship between the dynamic pressure, or stress, developed against the mill shell and the mean fracture stress of the material being ground, and the relationship between temperature and grinding efficiency. Although there are conflicting opinions about the relative importance of impactive versus abrasive grinding, it is well known that skilled operators of tube mills can determine whether such a mill is properly 'on' or 'going off the grind* by listening to the sounds made as the pebbles knock against the shell liners. The aggressive environment in an industrial mill and its rotary motion are the greatest obstacles to the insertion of sensors inside the mil! and the continuous recording of information from them. It is suggested that the bolts used to clamp the liners and lifter bars to the shell of a mill can be instrumented in a variety of ways so that important information can be obtained from inside the mill relating to its load-grinding action. This information can be accessed telemetrically by means of a transmitter equipped with an aerial and mounted on the mill shell. Results obtained with instrumented bolts that had been fitted to the no. 5 tube mill at East Driefontein Gold Mine are discussed.

144 SESSION C9 Instrumental Analysis

Paper G9/1

MODERN TRENDS IN X-RAY POWDER DIFFRACTION

by

R.L. Snyder*, H.E. Gobelf, and G. Hubert! •Alfred University, U.S.A. t Siemens AG, West Germany A number of events have occurred in recent years which have led to a resurgence of 1 interest in the techniques and applications of X-ray powder diffraction. The paper ! surveys these developments with an emphasis on the latest trends and their real and ! potential applications. The discussion is divided into three areas: instrumentation, I software, and applications.

Instrumentation The automation of the conventional 6-29 diffractometer with a diffracted-beam monochromator and scintillation detector during the 1970s led to a number of advances in the field of materials science. Modern developments in 0-9 and Seeman-Bohlin geometries, particularly with the use of incident-beam monochromators and position- sensitive detectors (PSD), have brought a host of previously inaccessible problems under the scrutiny of the powder diffractionist. These techniques, particularly when coupled with high-temperature and high-pressure cells and the very high intensities available from synchrotron and rotating anode X-ray sources, hold the promise of solving some of the most intractable problems of the chemical and physical changes induced by subtle- phase changes and non-stoichiometries. Recent advances in resolution and accuracy, along with time resolution (with PSDs), have already shown their power in giving new insights into reaction mechanisms. The use of Si(Li) detectors with white synchrotron sources also offers the possibility of exciting new applications of energy-dispersive diffraction.

Software Modern trends in both application and control software reflect both the hardware developments, which make ever more powerful computers available in the laboratory, and the system developments, which have produced sophisticated multi-user time-sharing systems for minicomputers. Control algorithms have brought full intelligence to all aspects of data collection except for the most common pattern scan. Peak-finding techniques have not improved significantly since the second-derivative procedures used in the initial systems. However, the use of interactive graphics, particularly on colour- display terminals, led recently to a great improvement in data evaluation. Recent developments in sample preparation (i.e. spray drying) and in data processing have made significant advances in quantitative analysis.

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\ A large amount of current research is focused on developing general methods for ! profile-fitting diffraction peaks. Investigators have already achieved powder patterns i , with figures-of-merit (Ffj) in excess of 200 with average A IB as low as 0,001°. These j new levels of accuracy have created the possibility of a wide range of new applications. ! : Applications Phase identification by computer using both first-generation (Johnson-Vand) and second-generation (Hanawalt) procedures have shown considerable success with even the most complex patterns. Current trends are in two directions. The first is the development of high-quality user files of the patterns most frequently encountered in a particular laboratory. With the accuracies obtainable through high-resolution instruments and profile fitting, the probability of the completely correct identification of complex mixtures via the Hanawalt type of search is greatly enhanced. These high accuracies also lead to a high probability of success for modern computer-indexing procedures. Successful indexing leads to the possibility of phase identification from lattice parameters, independent of intensities, via the CRYSTAL DATA data base. Profile- ! fitting deconvolution techniques, which extract the specimen-broadening function, hold '.„ the possibility of separating individual phases from a complex pattern for individual identification. A great amount of activity is directed at the whole-pattern profile-fitting procedure of Rietveld applied to the X-ray analysis of crystal structures. It is believed that the next few years will bring these methods to bear on the most difficult problems of materials science and solid-state chemistry. The latest developments in quantitative analysis have greatly increased the precision of this method. These results should lead to an increase in the industrial real-time (on- stream) control of production processes. The very high signal-to-noise ratios available from modern automated instrumentation have already been shown to lower the limit of phase detectability into the parts per thousand to parts per million range. The near future should see even structure analysis extended to crystalline grain-boundary phases. The advent of the linear position-sensitive detector has opened a wide range of time- resolved studies. One is now able to follow, not only the kinetics of a solid-state reaction, but also the phase and structural changes which occur during the application of any external stimulus such as heat, pressure, and magnetic or electric fields.

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Paper C9/2

AUTOMATIC ON-LINE ANALYSIS WITH ATOMIC-ABSORPTION SPECTROPHOTOMETRY

by

R.V.D. Robert and G.T.W. Ormrod Council for Mineral Technology South Africa A survey of the literature reveals surprisingly few on-line applications using atomic- absorption spectrophotometry (AAS), and automation of the technique is usually 7- confined to laboratory use. One of the major reasons for this is possibly the reluctance of potential users to leave j the flame of an AAS instrument burning unattended. Modern instrumentation, however, ' has many built-in safety features so that this is no longer a disadvantage. The 'tell-tale' j gold analyser developed by Brandt el al. uses flame AAS, after preconcentration by solvent extraction, for the measurement of the gold in the tailing solutions resulting _ from the cyanidation of gold. A number of these analysers are operating safely on several gold-extraction plants. It is predicted that the increased employment and acceptance of AAS with electrothermal atomization (ETA), as opposed to flame atomization, will result in a greater use of AAS in automatic on-line analysis. The direct sequential determination of up to six trace elements in flowing streams of cooling water from power plants is one example. The Council for Mineral Technology (Mintek) has developed and tested an automatic on-line gold analyser involving AAS and ETA, which is primarily for use on carbon- in-pulp extraction plants. Such an analyser installed on a gold-extraction plant permits the measurement of gold in solution at concentrations as low as 0,002 mg/1. The method is precise and accurate, and requires only a few microlitres of solution per measurement. Sample solutions taken from two or three of the six vessels on the plant are measured for their gold content, and the results permit reliable predictions to be made of the progress of the adsorption process. The sampling system consists of three polythene sample probes that are permanently positioned so that their lower halves are immersed in the pulp contained in vessels 4, 5, and 6 of the plant. Each probe has a porous polythene filter candle at its lower end through which the cyanide solution containing dissolved gold is filtered. A flow of nitrogen gas controlled by a solenoid valve lifts the filtrate, which is conveyed to the , analysis room via plastic tubing. In the analysis room, the solution flows into a vessel i that retains about 50 ml of the solution and allows the remainder to flow to waste. Another J solenoid valve allows the transfer of the retained solution to a sampling cup in which 0,5 ml is retained for measurement. The solenoid valves controlling the sampling are activated by a programmable sequence controller that also controls the operation of the AAS and furnace atomizer. A Varian 1275 Atomic Absorption Spectrophotometer and a Varian GTA 95 furnace system are used in the measurement of the gold concentration in the three sample

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> solutions and in a calibration standard solution. Between 5 and 20/*1 of the solutions, J together with lOfil of a nickel matrix modifier, are transferred in turn to a graphite { j furnace, where they are dried, ashed, and atomized. The gold concentration is then j ] measured relative to the calibration standard. } The analysis of the three samples and the standard solution is performed automatically ! at regular chosen intervals. The operating costs are limited to those for nitrogen gas and graphite atomization tubes. The results of an extensive test on the operation of this analyser are presented.

Paper C9/3

IMAGE ANALYSIS OF GOLD-BEARING ORES AND PRODUCTS

by t j E.J. Oosthuyzen ~~ Council for Mineral Technology ; South Africa There is an increasing need for the mineralogical investigation of gold in ores and products that contain about 1 p.p.m. (by mass) or less of gold. Since microscopic ; measurements on polished sections relate to volume fractions, the relative proportion of gold, because of its high relative density, is diminished by a factor of about six. Therefore, very large samples are needed if the microscopic observations are to be statistically adequate and the measurements realistic. For example, the exposed area of a particle of gold with a diameter of about 50 pm when sectioned accounts for 1 p.p.m. (by mass) of gold in 25 standard-size polished sections; similarly, in one polished section, 4 gold grains with diameters of about 5 jtm would represent the same amount of gold. The use of conventional manual microscopic methods is therefore unsuitable. However, with an automatic image analyser to search for and to measure the gold particles, the mineralogical investigation of gold in this type of material becomes feasible. This paper describes a procedure that involves the sampling of gold-containing ores and products, the search for gold in a sample, and the determination of the minerplogical parameters. The image analyser, which is programmed to search automatically for gold particles, makes use of the high reflectivity of gold in the wavelength region above 520 nm. The entire polished section is scanned and, when a gold particle is found, the coordinates of the microscopic field are recorded and stored. On completion of the automatic scan, all the microscopic fields where gold has been found are displayed sequentially for verification by manual methods. The mineralogical determinations include estimates of the amount of gold present and various measurements of the size of the gold grains, their mineral associations, and their shape. Other aspects discussed in the paper include the sampling procedure on a macroscopic as well as on a microscopic scale, the preparation of samples, and

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statistical evaluation in sampling and measurement to establish a norm for adequate sampling. The method was applied to a study of samples from the Klipwal Gold Mine, as well as to a study of dump material with a head value of 1 p.p.m. of gold and about 3 per cent pyrite. In the latter study, the method was applied to samples of the bulk material, separated pyrite fractions, and gold particles recovered from the pyrite by dissolution and centrifuging. The results are discussed and evaluated against results obtained by chemical analysis.

Paper G9/4

THE APPLICATION OF X-RAY FLUORESCENCE SPECTROMETRY IN MINERALS PROCESSING

by

J.P. Willis University of Cape Town South Africa X-ray fluorescence spectrometry (XRFS) is a non-destructive instrumental analytical technique which has been applied in minerals processing for approximately the last twenty years. In industrial use, XRFS is capable of determining all elements with Z> 11 (Na) at the concentration levels of major, minor, and trace elements. XRFS cannot be used to determine important elements such as hydrogen, lithium, beryllium, or boron, and the determination of carbon, oxygen, and fluorine is extremely difficult, if not impossible, under industrial conditions. Solids, liquids, and powders can be analysed by XRFS, and, being non-destructive, it is an exceptionally useful method for rapid qualitative analysis. XRFS can be divided into two major types: energy-dispersive X-ray spectrometry (EDXS), and wavelength-dispersive X-ray spectrometry (WDXS). Using EDXS, a qualitative analysis can be made in as little as 2 seconds for major elements, while 60 seconds is sufficient to determine all the major, minor, and trace elements at 10 to 19 parts per million. A corresponding analysis by WDXS could take as long as 60 minutes. The lower limits of detection with XRFS are in the range 2 to 20 p. p.m., and depend on the atomic number of the analyte element, the sample composition, the counting time, and the instrumental conditions. XRFS is strongly susceptible to matrix effects, which can be very severe. However, since these effects are both predictable and correctable, timeous steps can be taken to eliminate any problems that might arise, and XRFS can produce extremely accurate and precise analytical results. Although XRFS is regarded by some as the single answer to one's analytical problems, there is, unfortunately, no such method available today, and XRFS should be regarded

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as but one of many techniques in the analyst's armoury. It is particularly complementary to techniques such as neutron-activation analysis, atomic-absorption spectrophotometry, and optical-emission spectrometry (OES), especially when the last-named is used with an inductively coupled plasma (ICP) source. In fact, probably the most competitive analytical technique today to XRFS is OES-IGP, but for this method the sample must be in solution, which in many instances is a serious drawback and can limit its use. In industrial applications, where speed is essential and the main interest is usually in the concentrations of the major and minor elements, it is common to find multichannel XRFS instruments in general use. With modern instrumentation, it is possible to use up to twenty-eight different wavelength-dispersive channels simultaneously. If one channel port is used for an energy-dispersive spectrometer, this considerably enhances both the element coverage and the versatility of the instrument. The total analysis time is that for the element or channel with the longest counting time. With multichannel XRFS it is possible to analyse hundreds of samples per day, and often the speed of sample preparation is the limiting factor on throughput. In a research laboratory, as opposed to the production floor, a single-channel wavelength-dispersive spectrometer is generally preferred. XRFS is commonly applied to the analysis of ore-dressing products, slags, solutions, slurries, glass, ceramics, metals (both ferrous and non-ferrous), cement, ores, and minerals. The list is endless. Owing to its speed, non-destructive properties, and ease of automation, XRFS is frequently used in production control. For example, by determining element concentrations in situ in production feed and tailings lines, and using feedback control, production parameters can be automatically changed to maintain the product within specification. Various examples of XRFS applications in the minerals-processing field are given and discussed, together with modern developments.

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SESSION CIO ' Instrumental Analysis (continued) ; Paper C10/1

1 APPLICATION OF ION CHROMATOGRAPHY TO MINERAL PROCESSING j by C. Pohlandt Council for Mineral Technology South Africa The most significant contribution of ion chromatography (IC) to analytical chemistry is undoubtedly the separation and determination by conductometry of strong-acid anions. When IC was introduced in 1977, the separation of a mixture of seven anions took | approximately 20 minutes. With the highly efficient columns currently available, such I a separation can be conducted in 6 to 8 minutes. The apparatus employed in IC consists of a pump, an injection loop, a column, and a detector. Although similar to the apparatus conventionally used in high-performance liquid chromatography, IC equipment differs I in that it is essentially inert to the corrosive eluants that are usually necessary for the """" displacement of analyte ions from pellicular ion-exchange resins. Also, it makes use of a second column or suppression device, the sole purpose of which is to neutralize the eluant so that the analyte ions can be measured conductometrically. Liquid chromatographic techniques require a sample to be in solution, and in IC aqueous solutions are preferred since this permits, with little more than a dilution step, the determination of a large variety of ionic species in solutions and effluents such as those resulting from hydrometallurgical processes. Particularly noteworthy in this respect is the determination of cyanides in solutions encountered in the processing of gold. Cyanide is normally converted to weakly ionized hydrocyanic acid in the suppressor, and therefore cannot be determined conductometrically. By omitting the suppressor and using amperometric detection, one obtains a reaction between the cyanide ions and the silver electrode employed. The current resulting from oxidation of the electrode is then proportional to the concentration of cyanide in the solution. IC coupled with amperometric detection thus allows the determination of highly toxic ionizable cyanides with great rapidity (2 minutes), precision (a relative standard deviation of 0,112), and sensitivity (a limit of determination of approximately 10/*g/l for cyanide). Most methods currently used for the determination of the 'total' amount of cyanide present in a sample involve conversion of the cyanide species to hydrogen cyanide by acid treatment under reflux, and a subsequent distillation step. These procedures are slow and subject to interferences owing to the severe conditions required for the decomposition of stable metal-cyanide complexes. In an investigation undertaken af the Council for Mineral Technology (Mintek), it was shown that such complexes can be successfully ionized by irradiation with ultraviolet light, and that the combination of this method with IC gives a fast, reliable procedure for the determination of 'total* cyanide. In addition, it was found that individual metal-cyanide complexes can be successfully separated on columns with hydrophobic surfaces by eluants that have been modified with organic reagents.

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The application of IC to the analysis of solid samples is more difficult. Dissolution in acids is often not possible unless the analyte anion is present in sufficient quantities to allow a subsequent large dilution to be made. Specialized techniques, such as oxygen- flask combustion, have been used in the characterization of resins used in hydrometallurgical processes, and sample preparation by Parr-bomb combustion has been applied to the analysis of coal and oil shale. The determination of halogens in silicate rocks and other geological samples was made possible after combustion of the samples in a Leco induction furnace. Dissolution of samples via an alkaline-fusion step has been successfully applied to the determination of fluoride and chloride in tungstic oxide flotation concentrates, apatites, and fluorspars, and alkaline scrubber solutions of furnace gases have been successfully analysed for anions, including nitrite and nitrate. The capability of IC includes the determination of cations such as alkali metals, alkaline earths, and ammonium and transition-metal species. Because of the sophisticated techniques for metal analysis that are offered in most metallurgical laboratories, little effort has been made by workers to exploit IC in this respect. However, IC complements many other techniques, and valuable contributions can be expected from it in the analysis of cations, for example in the determination of rare-earth metals.

Paper C10/2 THE DIRECT ANALYSIS OF SOLID MATERIALS IN THE FORM OF SLURRIES BY THE INDUCTIVELY COUPLED PLASMA TECHNIQUE by A.E. Watson and P. Humphries-Cuff Council for Mineral Technology South Africa In conventional spectrometry with inductively coupled plasma (ICP), the sample, in the form of a solution, is introduced into the plasma discharge as an aerosol generated by some form of nebulizer. As a consequence of this, solid samples have to be dissolved prior to the analysis. Besides the extra time required for such a procedure, there are disadvantages (such as the dilution factor) inherent in a solution technique, including the fact that certain materials are extremely difficult, if not impossible, to dissolve. It is clear that, if the solid material could be introduced direct into the plasma without prior dissolution, there would be a considerable saving in the time normally spent on sample preparation, as well as a significant improvement in the lower limits of determination because of the lower dilution factors. A procedure was developed that consists in the nebulization of a suspension of the powdered sample in water, or other appropriate solvent, by means of a V-type, high- solids nebulizer and the transfer of the aerosol direct to the plasma. The technique was used in the analysis of activated carbon from the carbon-in-pulp process for the recovery of gold from cyanide solutions, as well as of resin from the resin-in-pulp process. . The limits of detection, the precision, and the accuracy of the technique are discussed, and results of the analyses of various materials are presented.

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